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J^ 8891 



Bureau of Mines Information Circular/1982 



Premining Investigations 
for Hardrock Mines 

Proceedings: Bureau of Mines Technology 
Transfer Seminar, Denver, Colo., 
Sept. 25, 1981 



Compiled by Staff, Bureau of Mines 




UNITED STATES DEPARTMENT OF THE INTERIOR 



^^.jca J^^^f^^. ^^-^^T^/-"^^--^) 



Information Circular 8891 

^ 



Premining Investigations 
for Hardrock Mines 

Proceedings: Bureau of Mines Technology 
Transfer Seminar, Denver, Colo., 
Sept. 25, 1981 



Compiled by Staff, Bureau of Mines 




UNITED STATES DEPARTMENT OF THE INTERIOR 
James G. Watt, Secretary 

BUREAU OF MINES 
Robert C. Horton, Director 




This publication has been cataloged as follows: 






.01 



Bureau of Mines Technology Transfer Seminars (1981 : 
Denver, Colo,) 

Premiiiing investigations for hardrock mines. 

(Information circular/ U.S. Dept. of the Interior, Bureau of Mines ; 
8891) 

Includes bibliographical references. 

Supt. of Docs, no.: 128.27:8891. 

1. Mine examination— Congresses. I. L'nited States. Bureau of 
Mines. II. Title. III. Scries; Information circular (United States. 
Bureau of Mines) ; 8891. 

TN295.1J4 rTN272l 622s (622'. 14) 82-600057 AACR2 



PREFACE 

This Information Circular summarizes the results of recent Bureau of 
Mines research concerning improved methods and techniques employed dur- 
ing premining research for metal and nonmetal mining. The papers are 
"Vi only a sample of the Bureau's total effort in the areas of ground con- 

trol, maximum resource recovery, and efficient extraction technology, 
but they delineate the major advances in the area of premining research. 
Much of the technology discussed has been refined from previously used 
techniques and is applicable to other areas of mining and to other in- 
dustries such as petroleum. 

The technical presentations reproduced herein were made by Bureau 
technical personnel of the Technology Transfer Seminar on Premining 
Investigations for Hardrock Mines given September 25, 1981, in Denver, 
Colo. Those desiring more information on the Bureau's research in the 
areas described or other general information should contact the Bureau 
of Mines Technology Transfer Group, 2401 E Street, N.W., Washington, 
D.C. 20241, or the appropriate author. 



V 



CONTENTS 



iii 



Page 



Preface 1 

Abstract 1 

Introduction, by Joseph L. Condon 2 

High-Resolution Seismic Methods for Hard-Rock Mining, by Frank Ruskey 4 

Application of the Electrical Resistivity Method to Mining Problems, by 

Richard G. Burdick 29 

Electromagnetic Ground Radar Methods, by Richard J. Leckenby 36 

In Situ Neutron Activation Analysis, by George J. Schneider 46 

Development of An In-Hole Replaceable Diamond Core Bit System, by W. C. Larson, 

W. W. Svendsen, R. E. Cozad, and J. R. Hoffmeister 55 

Structural Design for Deep Shafts in Hard Rocks, by Michael J. Beus and 

Samuel S. M. Chan 65 

Borehole Deviation Control, by E. H. Skinner and N. P. Callas 79 



PREMINING INVESTIGATIONS FOR HARDROCK MINES 

Proceedings: Bureau of Mines Technology Transfer Seminar, 
Denver, Colo., September 25, 1981 

Compiled by Staff, Bureau of Mines 



ABSTRACT 

These proceedings consist of papers presented at a Bureau of Mines 
Technology Transfer Seminar in September 1981 for the purpose of dis- 
seminating recent advances in mining technology in the area of premining 
research. The introduction and descriptive papers discuss techniques 
and instrumentation used in premining research for metal and nonmetal 
mining and shaft design and borehole control for premlne planning. 



INTRODUCTION 
By Joseph L, Condon 1 



The Bureau of Mines traditionally de- 
fines premining investigations as the 
series of studies that follow an explora- 
tion discovery and precede production 
from a mineral deposit. In the mining 
industry the studies may be the responsi- 
bility of a specialized group within a 
con5)any, or they may be shared between 
exploration and production organizations. 
Typically, the studies at one prospect 
are sequential, with the start of a suc- 
cessive study dependent on the successful 
conclusion of a prior study, until both 
economic and engineering feasibility of 
production from the deposit are demon- 
strated. Beyond that point in time, more 
information is necessary to describe the 
characteristics of the ore body for de- 
velopment and detailed mine planning. 
After mining has started many techniques 
developed for premining investigations 
are usable to define the ore body beyond 
the working face for production planning. 

The Bureau of Mines research in premin- 
ing investigations is conducted under the 
Assistant Director — I4ining Research. It 
includes the development of methodologies 
for collecting and analyzing information 
on ore body characteristics to enhance 
mine production and safety. The aim of 
the Bureau research program is not the 
development of technology for its own 
sake, but to develop innovative methods 
that are better and cheaper than existing 
techniques to replace and supplement cur- 
rent industry practice and to improve 

^Research supervisor, Denver Research 
Center, Bureau of Mines, Denver, Colo. 



conventional techniques to increase their 
capabilities and lower their costs. 

The Technology Transfer Seminar on Pre- 
mining Investigations for Hardrock Mines 
describes and demonstrates the results 
of recent Bureau research. Pulsed and 
continuous wave ground radar systems that 
are capable of mapping ore body charac- 
teristics with high resolution have been 
shown feasible for application from the 
surface, in boreholes, and from the work- 
ing face underground. High-resolution 
seismic methods are now available that 
are capable of utilizing higher frequen- 
cies and obtaining geologic information 
from much shallower depths conpared with 
the conventional seismic reflection tech- 
nology for the petroleum industry. 
Experience with coal strata has shown 
that the high-resolution seismic reflec- 
tion method can map subsurface features 
in media unfavorable for the propagation 
of electromagnetic radiation, while the 
ground radar method can provide superior 
details within the coalbed. Analogous 
application of the two techniques is an- 
ticipated in hard-rock mining. 

A computer program has been developed 
to map lineaments such as geologic faults 
and fractures from satellite imagery. 
Past research has demonstrated that 
ground control problems are often present 
in underground mines below the intersec- 
tion of lineaments that are apparent in 
remote sensing data. Readily available 
and relatively low-cost resistivity 
equipment is usable for fast reconnais- 
sance to verify remote sensing data. 



An in situ neutron activation analysis 
system has been developed to assay ore in 
place from small-diameter access bore- 
holes. The logging sonde in the sys- 
tem has a 4,000-channel, microprocessor- 
controlled analyzer with digital communi- 
cation over a standard logging cable to 
the surface. The system permits a 
laboratory-quality gamma ray spectrometry 
from several thousand feet deep in the 
earth. 

A viable diamond core bit system that 
can be replaced inside the drill hole has 
been developed to significantly improve 
the efficiency of core drilling by mini- 
mizing the time and labor required for 
bit replacement. Techniques for con- 
trolling the deviation of drill holes 
have also been developed. Even with 
the availability of in situ assaying 



technology, core drilling will be re- 
quired for "hands on" samples and for 
that reason alone will never be totally 
replaced. 

Finally, a method to design deep shafts 
in hard rock will be discussed. The 
design technique relies partly on geo- 
technical data that can be obtained with 
existing instruments and mechanisms. 

The results of Bureau research in the 
subject area are impressive. With the 
new techniques available, the mining 
industry should improve its practices for 
premining investigations. The Bureau of 
Mines continues to provide real solutions 
to meet the challenge of varying geologic 
conditions and changing mining tech- 
nologies for various types of mineral 
deposits. 



HIGH-RESOLUTION SEISMIC METHODS FOR HARD-ROCK MINING 
By Frank Ruskey^ 



ABSTRACT 



Seismic procedures for applications in 
mining problems are continually improv- 
ing. Increasing numbers of mining com- 
panies are using the technology or ex- 
ploring its possibilities. This paper, 
primarily intended for the casual user, 



describes techniques and procedures, con- 
straints, and pitfalls that may be en- 
countered. Results of tests at various 
sites are presented to show the possibil- 
ities of the technology. 



INTRODUCTION 



This report provides the mining indus- 
try with the results of research and 
evaluation performed by the Bureau of 
Mines in high-resolution seismics for 
mining applications. The work arose out 
of a need to find and develop a geophysi- 
cal technique that could locate channel 
sands and faults from the surface for 
mine safety planning. From an evaluation 
of various geophysical methods, magnet- 
ics, electrical, gravity, and seismic, 
the seismic method was soon considered 
the best tool for locating geologic fea- 
tures related to mineral deposits, such 
as faults and channel sands. The other 
geophysical tools have their unique value 
for other mining applications. 

Once the value of seismics was estab- 
lished, it became apparent to Bureau of 
Mines researchers that a multifold 
approach to the problem was necessary. 
In-house research was considered essen- 
tial to develop the concepts towards 
achievable goals that could be used by 
the mining industry. Coupled with this 
was a need to stimulate potential users 
into realizing the value of seismics for 
improving safety and providing extraction 
economies. It was believed that the min- 
ing industry would be interested in the 
viability of using seismics for solving 
mining problems , and that geophysical 
service companies could be made aware 
that a potential market exists for their 
services. Accordingly, contracts were 



Geophysicist, Denver Research Center, 



Bureau of Mines, Denver, Colo. 



let over several years to test feasibil- 
ity, configure an optimum geophysical 
data acquisition system from off-shelf 
equipment, and perform tests in many min- 
ing environments and for a variety of 
mining problems. Commercial seismic pro- 
cessing centers were contracted to ana- 
lyze data, using their already developed 
oil exploration expertise. They were 
chartered to test a variety of tech- 
niques, to determine which were best for 
the many possible types of mining prob- 
lems, and to simplify each technique so 
that a practical, economical service can 
be engendered. In addition, Bureau per- 
sonnel have presented the results of 
their work at many conferences and sym- 
posiums to let the mining industry know 
that, indeed, here is a tool that can 
save both lives and, potentially, milli- 
ons of dollars in production costs. 

Also, it has been recognized that there 
are many problems associated with mining 
other than fault and channel sand deline- 
ation that can be aided with applications 
of seismic technology. Among these are 
abandoned mine location, ore zone bounda- 
ries, split coal seams, vein delineation, 
and others. Some of these have been the 
subject of preliminary investigations by 
Bureau researchers. Some of these re- 
sults are described herein to indicate 
the full potential of the seismic tech- 
nique. Although the field examples pre- 
sented in this report were taken from 
tests in coal, the procedures are identi- 
cal with those used for any ore deposit 
in any geological environment. 



The report includes, in addition to 
Bureau work, that of some Bureau contrac- 
tors over the past 5 years. Their pio- 
neering efforts have been significant in 



bringing seismics for mining applications 
to its present usefulness. Additionally, 
Sheriff (J_3)2 presented a brief descrip- 
tion of commonly used seismic terms. 



RECOMMENDED SPECIFICATION FOR DATA COLLECTION AND RECORDING SYSTEM 



This specification is based upon the 
field experience of Bureau of Mines per- 
sonnel over the past 5 years. It is in- 
tended as a guide to potential users or 
purchasers in their shopping, or as a 
basis for evaluating proposing seismic 
service con^janies that may be under con- 
sideration to perform work for their min- 
ing con5)any. While it is not intended to 
lock out potential service companies that 
would be able to solve many problems and 
achieve useful results using less sophis- 
ticated equipment, the recommendation is 



that a higher ratio of success can be 
anticipated, particularly for knotty 
problems or terrain conditions, if the 
following minimal criteria are met. 

Data Acquisition — ^Minimal Specifications 

1. Digital System: A digital system 
is a must. Data sample rates of 1 mil 
are acceptable for most applications. 
Sample rates of 0.5 mil or 0.25 mil will 
aid higher resolution. This is shown 
cursorily in table 1. 



TABLE 1. - Table of recordable frequencies at various sample rates 



Sample rate, 
mil 


Frequency , 
Hz 


Approximate 

resolution, 

ft 


Approximate depth 

of applicability, 

ft 


2 

1 
.5 
.25 
.125 


128 

256 

512 

1,024 

2,048 


4 
2 
1 
1 
.5 


1,000-10,000 

300- 3,000 

100- 500 

200- 300 

50- 200 



2. Stacking Capability: For most ap- 
plications, the data acquisition system 
should be able to stack up to 20 repeat 
shots, if necessary. An acceptable sys- 
tem, but cumbersome, would be one where 
each shot or stack element is recorded on 
tape and stacked at the processing 
center. 

3. Visual Recording for Field Monitor- 
ing: The field operator should be able 
to look at his field data and make judg- 
ments on whether to alter his field pro- 
cedure, for optimal signal returns, as 
subsurface conditions change. Simultane- 
ous visual recording of all traces is 
desirable. Trace by trace recording is 
acceptable, but slow and cumbersome. 

4. Amplifier Gain: Each recording 
channel should have 96-db gain, in 
approximately 12-db steps, to enable 
sufficient gain over a full spread, and 



to taper the spread gains 
to far geophone positions. 



from near-shot 



5. Record Time: The system should 
record a minimum of 0.5 sec at the 0.5- 
mil sample rate. 

6. Gain Ranging: This feature (some- 
times called automatic gain control) can 
be offset for shallow work, if sufficient 
initial energy can be put into the 
ground. If low-level inputs are used, or 
much stacking is required, then gain 
ranging is desirable. 

7. Magnetic Tape: The data must be 
digitally recorded for later conputer 
processing. The subtleties of shallow 
seismic analysis can only be brought out 

^Underlined numbers in parentheses re- 
fer to items in the lists of references 
at the end of each paper. 



with computer processing. Computer- 
coii^)atible reels are most convenient for 
mass storage volume and for later ease of 
mounting on data processing equipment, 

8. Seismic Detectors: Geophones (ve- 
locity detectors) or seismic accelerom- 
eters are equally good. Geophones have 
proven designs, ruggedability , reasonable 
response, low cost, and minimal mainten- 
ance. Accelerometers require line ampli- 
fiers that need periodic maintenance or 
battery changing. The use of subarrays 
will compromise resolution to a degree, 
but they provide greater field versatil- 
ity for improved signal-to-noise capabil- 
ity over single geophone layouts. 

Alternate Systems 

Although this report makes a strong 
case for sophisticated data acquisition 
systems, it should be kept in mind that 
many of the simpler systems now on the 
market can yield significant data for 
problem solving. The use of such systems 
is recommended where budget, terrain, or 
personnel constraints inhibit the use or 
acquisition of the more expensive gear. 

Additionally, systems are being devel- 
oped that provide almost immediate pro- 
cessing of the filed data. Such systems 
are cost effective and will provide a 
useful tool to the coal company that 
wishes to perform its own surveys at min- 
imal cost. One such economic system is 
noteworthy of description. 

As part of their Coal Geophysics Re- 
search Project, the U.S. Geological Sur- 
vey is developing a small-crew, rela- 
tively inexpensive shallow target seismic 
system. Figure 1 shows the four prototye 
elements of the recording system; a 12- 
channel, signal-enhancement, 10-bit seis- 
mograph, with sample rates from 0.05 to 
2.0 msec; an interface box containing 
filters and signal-switching units; a 12- 
channel monitoring oscillograph; and a 
digital tape recorder. A special one- 
wheel carrier is used to first move a 
posthole-type drill to the site, and then 




FIGURE 1. - Inexpensive, alUpurpose shallow 

seismic system. (Courlc .sy, U.S. 

(reohxjical Survey.) 

bring in the recording equipment. With 
its handles turned down, this carrier 
becomes the covered table upon which all 
elements of the data acquisition system 
are mounted. 

Data are processed using the desk-top 
computer system shown in figure 2. This 
equipment is carried in a small trailer. 
By having the conqjuter system located 
near the center of field operations, it 
is possible, on the evening of the day on 
which the data were taken, to review 
their quality and to make an initial 
evaluation of results. 




FIGURE 2, - Data Processing system for U.S. Geological Survey seismic system. 

(Cour/r sy, U.S. Ge oloyical Surney.) 

FIELD TECHNIQUES 



This section provides a partial de- 
scription of field techniques used by 
Bureau personnel with reasonable success. 
It also includes some information on pro- 
cedures used by others, primarily commer- 
cial crews, that are known to be success- 
ful. The coverage of all procedures is 
intended as a user's guide only and 
recognizes that there may be many other 
successful techniques under wraps, or as 
yet unproven. Greatest success will be 
achieved by operators that have the ver- 
satility to use several procedures, in 
order to tackle problems where geologic 
conditions vary. Optimum field tech- 
niques are field condition dependent. 
Geologic conditions, terrain conditions, 
groundwater, and cultural conditions may 
each have to be approached with a differ- 
ent procedure. 

Geologic conditions include the lithol- 
ogy of the deposit and the overlying 
strata and structural conditions, such as 



faulting, fades changes, or unconform- 
ities. Each of these will affect seismic 
signal returns, and one should be at 
least generally aware of their likelihood 
in the survey area. 

Terrain conditions include trees 
and other vegetation; hills, valleys, 
streams, or other bodies of water; swampy 
versus dry land; and heavy, spongy soil 
mantle versus a thin mantle or exposed 
outcrop rock. Included here also are ac- 
cess roads, habitation, buildings, roads, 
and pipelines, etc. Each of these has 
its adverse effect on a proposed survey. 
However, each condition can be reasonably 
tackled with a versatile crew, provided 
they are allowed the time to tailor their 
approach to changing conditions. 

The following section is presented to 
indicate some of these problems and some 
potential solutions. 



SOME FIELD PROBLEMS AND POSSIBLE SOLUTIONS 



Trees and Other Vegetation 

Creates Spurious Noises If Windy 

Wind noises alone, or rocking trees and 
vegetation, are usually low frequency and 
random; solutions are to use (a) higher 
frequency (100 Hz) geophones and (b) re- 
peat shots with stacking to improve 
signal-to-noise response. 

Inhibits Access 

If vegetation cannot be bulldozed away, 
the geophone spreads can still be laid 
out by using off-end, fan, or 3-D shoot- 
ing techniques. 

Hills and Valleys 

Source of Spurious Reflections That Mask 
Wanted Seismic Signal Returns 

The spurious reflections usually show 
up as a strong band in the seismic sec- 
tion. Computer processing migration 
techniques are effective in subduing this 
source of noise. In extremely hilly 
terrain 3-D or concentric circle tech- 
niques described in later sections of the 
report may be used effectively. 



Changes in Lithology 

The effect of changes in lithology is 
shown hypothetically in figure 3, where 
it may be seen that the longer travel 
path from shot-point B to geophone B' 
reduces con5)atibility with the results 
from shot-point A. Hence, a confusing 
seismic section, as shown in figure 4, 
may ensue. Part of the solution is data- 
processing-center dependent, wherein 
traces with abrupt elevation changes, 
used for common depth point (CDP) stack- 
ing, are trace analyzed and adjusted. 
This may be a tedious manual process, but 
it is necessary for shallow reflection 
work. In the field, some amelioration 
can be achieved by placing geophones at 
nearly the same contour elevation — if 
possible, or by using the 3-D or concen- 
tric circle techniques. Static and move- 
out corrections have to be applied with 
discretion. 

Inhibits Access 

If the seismic crew is vehicle depen- 
dent with drills, recording truck, shoot- 
ing or energy source trucks, they will be 
limited to available access roads. For 
such cases 3-D or offset shooting can be 



I Geophones^ 




FIGURE 3. - Lithologic changes that may complicate seismic data returns. 



SHOTPOINT STATION 

CM^'»-i-,-^00000 
CMCMCMCMCMCMCVJCsJCMCMcM 




■^'^^^'t^i-:i:fii^^^^^dl 0.1 




PROCESSED FOR 



POLARITY 
REVERSED 



U.S. BUREAU OF MINES 



SHOT RY U.S BUREAU OF MINES p^rty — 

DATE 6/IQ/76 ENERGY SOURCE CAPS 

FILTER. - - _rt; |NSTR||MFMT<; INPUT/OUTPUT DHR 1632 

SAMPLE INTERVAL 1^2 ITU RECORD I Fiar.TM 2.0 ..r 

SHOT DEPTH Q-IO (, CHARGE SIZE ^_£AES____ 

RECORDING GAIN NONE SHOT POINT INTERVAL 50 It 



NUMBER OF GROUPS - 



-GROUP INTERVAL. 



SPREAD r.FnuFTRY 450 - 100 - ■)» - 100 - 450 



SAMPLE RATE l/g 

DD ANALOG TO DIGITAL CONVERSION 

niH REFORMAT 

nn RESAMPLE 

DD DEMULTIPLEX 

nS GAIN RECOVERY 

nS CDPSORT 

nn TRACE EDIT 

nifil VELOCITY ANALYSIS LOCATIONS FVFRY ?0 C DP'S 

nn MUTING 

START TIME NEAR OFFSET m> 

END TIME FAR OFFSET ms 

nS! TRACE BALANCE 

nn 0EC0NV0LUT:0N length mt % WHITENING 

DESIGN GATE i 

NEAROFFSET STARTTIME ms END TIME im 

FAR OFFSET STARTTIME m«. END TIME ml 

DESIGN GATE 2 

NEAR OFFSET START TIME til END TIME ml 

FAR OFFSET STARTTIME .-ns END TIME rm 

DESIGN GATE 3 

NEAR OFFSET START TIME miENDTIME mi 

FAR OFFSET STARTTIME msENDTIME mj 

(313 AUTO STATICS 

lain NMO CORRECTIONS 

ns DATUM CORRECTIONS DATUM J2QQ_tt Vl«. 

ICin COMMON DEPTH POINT STACK FOLO_i_ 

EUn FILTERING I FNCTH 126 



LOW CUT (Mil 


240 


0.0 


) mot. TIME 


200 



















nE COHERENCY STACK 

nn WAVE EQUATION MIGRATION 

nn 



400% COHERENCY STACK, 
polarity reversed 



SOUTH EAST^ 

FIGURE 4. - Marginal-quality seismic section due to underdeveloped field procedures. 



10 



considered. A fully portable system, 
such as developed by the USGS Coal 
Branch, previously described, would be 
useful for these conditions. 



natural land procedures are more 
state-of-the-art. 

Geophone Emplacement 



or less 



Streams or Other Bodies of Water 

Inhibits Access 

For most applications access is the 
problem. If the body of water is large, 
shallow marine procedures may be neces- 
sary. The stream or water body may 
restrict the laying out of seismic lines, 
or the placement of signal energy 
sources. Offset shooting or 3-D shooting 
are both well developed and provide prac- 
tical approaches. 



Swamps can be an asset or a detriment. 
Versatility of the field crew and their 
equipment repertoire is the solution. 
Hydrophones, rather than standard geo- 
phones, can now be used. If carefully 
placed into the water level of the swamp, 
their frequency response could be high. 
However, some swamps may act as a high 
frequency attenuator. Only experience 
and testing will tell. A problem arises 
in maintaining continuity between swampy 
and dry land data, if a mix of sensor 
equipment is used. 



Swampy Terrain Versus Dry Land 



Signal Sources 



Inhibits Access 

The solution to swampy conditions is 
similiar to that applied to streams or 
other small bodies of water. The dry 



Operating in a swamp environment neces- 
sitates the use of either explosives or 
sparkers. Here the need for a repertoire 
of signal sources is essential. 



USEFUL FREQUENCIES 



In high-resolution field surveys, 
the question of optimum source-receiver 
frequencies becomes an issue. High fre- 
quencies and high resolution are inter- 
dependent. Factors affecting choice of 
recordable frequency content are cost, 
speed of seismic productions, depth of 
penetration, and the magnitude of the 
geologic feature sought in the survey. 
For large features, such as large channel 
sands, faults with considerable throw, or 
the boundaries of an ore body, a lower 
frequency survey will be adequate. For 
such as ore veins, 
seams, small channel 
a higher frequency sur- 



small features 
splitting coal 
sands, or faults 
vey is necessary 



Equipment choices for high-frequency 
surveys become a factor. High- and low- 
frequency energy can be obtained, for 
instance, from explosives, controlled 
explosion chambers, and weight drops. 
Detector responses have to be com- 
patible with the anticipated frequency 
spectrum. Standard geophones, with a 
natural frequency above 40 Hz, are rec- 
ommended to help attenuate the inevitable 



low-frequency ground roll. Data acquisi- 
tion systems capable of acquiring the 
higher frequencies are essential. If 
digital, the sample rates shown in ta- 
ble 1 are important. If an analog sys- 
tem, response characteristics of the 
recording pen must be sufficient to re- 
cord the highest desired frequencies. 

For mining applications, frequencies of 
approximately 100 Hz are sufficient. 
Good returns at depths to 500 ft are 
readily obtainable, and resolution is 
adequate to provide meaningful interpre- 
tation of most mining problems. Broad- 
band data encompassing, perhaps, 80 to 
200 Hz have been found to be the most 
useful for mining applications. 

Definition of a given problem is not 
solely dependent upon resolution of the 
feature at the ore deposit. The seismic 
section has valid interpretation data 
from very near the surface to depths 
below the ore deposit itself. This is 
shown in figure 5, wherein a channel sand 
can be seen affecting the overlying sedi- 
ments, and faults (fig. 6) can be seen 



11 



OZl 






(:^Ai^^$^ 



iiifSI 

ill 



,r.^".^*>< 






ozi-i 



036- 



02 1.'!.- 



oze'i 



ilpliii 



: - - .-- 1 t 






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08S '31^11 




iii 



i^l^M' 



12 



throughout the entire geologic section. 
If the feature is sufficiently large to 
affect the overlying geology, then both 



it and its effect on the 
be detected and mapped. 



overburden can 



BURIED PHONES VERSUS SURFACE PHONES 



The question of whether to use buried 
phones, or merely surface phones, is an 
important consideration. 

buried phones means putting the geo- 
phones in drilled holes at the base of 
the weathering, or at the water table. 
These holes could range from 1 ft to 
lOU ft deep, with depths of 3 to 20 ft 
probably being nominal. The advantage of 
buried phones is that high-frequency and 
high-quality data can be obtained. Has- 
brouck ib) and Ziolkowski and Lerwill 
( 13 ) obtained frequencies of 100 Hz and 
produced excellent cross sections to 
depths of several hundred meters. The 
buried-phone scheme lends itself to the 
use of hydrophones with their excellent 
coupling characteristics in water. Also, 
single phones are used rather than sub- 
arrays; hence, phasing differences be- 
tween phones that invariably result in 
smearing of the high-frequency data are 
eliminated. The result is sharp fre- 
quency definition. Additionally, good 
production techniques can be worked out, 
wherein each geophone position becomes a 
drilled hole for energy-source explosive 
emplacement. Hence, each hole has a use- 
ful dual purpose, and little addition- 
al time is lost in drilling, planting 
phones, and shooting, etc. For critical 
or troublesome areas, where one wishes to 
obtain high detailing in the data, the 
use of buried phones may be a must. One 
should be prepared to use buried phones 
for critical areas. 

Bureau of Mines research, however, re- 
volves about using surface phones, some- 
times burying the phones just below 
the surface if it is windy. Occasion- 
ally, subarrays, with six phones in a 



series-parallel arrangement, are used to 
half -wave attenuate part of the ground- 
roll energy. For such applications, 
burying phones (3 ft or deeper) is too 
time consuming to consider. Although 
burying seems favorable at times, drill- 
ing holes for each phone can be diffi- 
cult. For instance, retrieving phones 
may be a concern should the weathering 
layer be particularly unstable and tend 
to bury them permanently. Further, Bu- 
reau researchers find that many mining 
applications, especially for abandoned 
workings in urban areas, are out of 
bounds for drilling and explosives use. 
Hence, the development of procedures for 
the use of repeatable surface energy 
sources has been given preference. 

Bureau researchers find that the seis- 
mic approach using repeatable signal 
sources has many distinct advantages over 
the use of explosives and buried geo- 
phones. Among them are repeatability of 
the source, the subsequent signal stack- 
ing, and the potential for weighted 
shots, all of which help increase signal- 
to-noise ratios. 

Many mining problems can readily be ad- 
dressed without the use of buried phones. 
Bureau experience suggests using surface 
techniques for most surveys, or the 
greater portion of any survey, to be fol- 
lowed up, perhaps in a few select trou- 
blesome areas where detailed answers are 
needed, by a few lines of buried phones. 
Once an answer has been found in one 
area, one finds that selective extrapola- 
tion to other areas of a cross section 
can be readily achieved to improve the 
total subsurface analysis. 



SINGLE GEOPHONES VERSUS SUBARRAYS 



Good field procedures necessitate ver- 
satility in using trade-offs to achieve 
"optimum" results. An important one 
is the use of single geophones versus 
subarrays. Both have a value, depending 



upon the problem being studied and on 
other circumstances, usually the seismic 
characteristics existing down to the tar- 
get area. 



13 



Mining applications are usually "shal- 
low seismic" applications and present a 
series of considerations. Among these is 
the travel path of the seismic energy 
wavefront. For shallow applications, 
this path is at a relatively obtuse angle 
(02) from the signal source, as shown in 
figure 7; while for deeper horizons, it 
is at an acute angle (0i). 

For single-phone applications, this 
merely means that the seismic energy im- 
pinges upon the phone at an angle, and 
the subsequent energy transfer to the 
coil is governed by the cosine of the 
angle from vertical, as shown in fig- 
ure 8. This, of course, means a loss of 
signal energy that may be overcome by 
repeat shots (stacking) or by increasing 
amplifier gains. Frequency response at 
the amplifier will, however, be exactly 
the frequency of the moving coil, which 
hopefully, if the geophone plant is good, 
will be the frequency of the incoming 
reflected wave. 

If an array of geophones is used, ei- 
ther to increase total energy returns for 
weak signals, or to provide ground-roll 
attenuation, then the angle of the upcom- 
ing signal may be significant. As shown 
in figure 9, the first phone in the array 
(A') receives the wavefront energy and 
begins to output signals to the amplifi- 
ers. Microseconds later, the second 
phone (B') receives its signal and out- 
puts signal to the same amplifier, 
slightly out of phase with the signal 
from the first geophone. This continues 
through the entire string of geophones 
and, of course, is accentuated the fur- 
ther the phones are separated from each 
other, or if the terrain is hilly. This 
effect is frequently called phasing or 
smearing. If high-resolution returns are 



being sought, this objective will be com- 
promised. For many applications, this 
may be satisfactory. For others, the 
important high-resolution signal nuances 
may be lost forever. 

Additionally, manufacturing tolerances 
for standard geophones are such that some 
variations in sensitivity occur (J)' 
While it should be understood that these 
tolerances are quite sufficient for the 
predominant oil exploration market, they 
may sometimes be marginal for high- 
resolution mining seismic applications. 
The effect of this difference in sensi- 
tivity is similar to that described 
above, in that if the units are in an 
array, then phasing and frequency smear- 
ing can occur. Sometimes it may be a 
problem — sometimes not. 

On the other hand, geophone subarrays 
are a powerful tool for attenuating un- 
wanted ground-roll signals. If well con- 
figured, subarrays may attenuate ground 
roll by as much as 24 db. Farr (_5) de- 
scribes this procedure. Because ground- 
roll returns invariably come in during 
the time window of wanted shallow reflec- 
tion signals, their attenuation to any 
degree is helpful. A strong combination 
can be achieved through repeatable sur- 
face sources (single or weighted), subar- 
rays, and carefully chosen offset dis- 
tances from source to geophones. The 
trade-off, however, will be some loss of 
resolution. 

In the Bureau's ongoing work, the pres- 
ent emphasis is to develop procedures 
around single-phone configurations with 
repeat signal inputs, and other window- 
ing procedures to eliminate ground-roll 
effects. 



AMPLIFIER CHANNELS 



Although the simplest of systems, hav- 
ing only one or two channels, can provide 
useful information for some problem ap- 
plications, the availability of at least 
12 channels is desirable; 24, 48, 96, and 
up are excellent. However, the more 
channels, the higher the initial costs, 
and, possibly, the production costs per 



mile of coverage. Also, when large num- 
bers of channels are used in linear sur- 
veys, offset distances to the far phones 
become a serious problem for shallow- 
mining applications. Bureau experi- 
ence for linear surveys indicates that a 
good tradeoff for a digital system is 24 
channels. For 3-D (three-dimensional) 



14 



Source 



Bo- 



Geophone 

n Surface 



Reflector I 




Surface 



Reflector 2 

FIGURE 7. - Seismic travel path considerations. 




Geophone 



FIGURE 8. - Geophone responseas a func- 
tion of ray path angle from vertical. 



/3,S,C, and D are potential 
signol energy at each 
geophone /I ', 6 ', C ', and D' 




To some amplifier 



Response of individuo 
geophones 



Composite response 
at amplifier of 

A',B',C',D' 




FIGURE 9. - Effect of wavefront impinging on a suborrcy. 



15 



applications, 48 to 96 channels or more 
are best. 

A 24-channel system allows for up to 
12-fold CDP (common depth point) stacking 
in a linear array, and more if 3-D grid 
shooting is performed. A 12-channel sys- 
tem providing 6-fold CDP stacking has 
been found to be minimum for mining 
applications. The 24-channel system, on 
the other hand, provides a greater versa- 
tility of operation. Triaxial arrange- 
ments can be configured, or single versus 
multiple phones used, without losing too 
much lineal progress of the line. Also, 
3-D grid shooting, which is showing prom- 
ise of being a powerful technique, can be 
implemented more readily. 

In all of this, however, one should be 
governed by the magnitude or complexity 



of the problem being studied. If a small 
project, such as locating a channel sand 
washout in front of active mine workings, 
is being tackled, a 12-channel or smaller 
system would be adequate. If it is 
desired to map an area of 1 square mile, 
it would be better to use a 24-channel or 
larger system. 

Some of the early work by Bureau re- 
searchers was with an 8-channel system. 
It was found to be quite sufficient for 
many applications, but limiting, at 
times, for performing large-scale or mul- 
tiphone studies. Figures 10 and 11 show 
some of this 8-channel work. The quality 
is reasonable, but additional coverage, 
with about the same amount of effort, 
could have been obtained with a larger 
system. 



INITIAL TESTS — NOISE SPREADS 



Of paramount importance in any survey 
is the need to perform a series of ini- 
tial tests to seismically characterize 
the site. This should minimally include 
an uphole survey in the site area so that 
the representative velocity distribution 
can be determined early in the survey to 
aid site characterization. Secondly, the 
area should be thoroughly evaluated, on a 
selective basis, to determine which of a 
variety of seismic approaches should be 
used. 

The velocity survey is important, be- 
cause seismic systems essentially measure 
time of wave front travel. Hence, in 
order to determine depths, which is what 
one really wants, it is necessary to know 
the wave velocity of the travel path. 
The ultimate accuracy of any survey will 
depend upon how well the velocity distri- 
bution is known. For much work, a good 
"average" velocity distribution will 
yield many solutions. This average 
velocity can be obtained from an uphole 
survey in a nearby drill hole, or from a 
refraction survey over the area, or from 
computer analysis of mo veout-t ravel path 
differences. Of these, the velocity up- 
hole survey is the best, although it 
represents only one region in the survey 



area. The others provide broad coverage 
but are heavily statistics dependent. 
Procedures for conducting these surveys, 
and the subsequent analysis, are thor- 
oughly covered in the literature, for 
example Dix (2^) and Embree (4^). 

Initial tests are so important that 
Bureau researchers unequivocally adhere 
to the dictum, "Do not proceed with the 
production survey until you have thor- 
oughly evaluated the site and its optimum 
field procedure requirement." The 
sought-for goal is the best possible 
signal-to-noise ratio in the depth zone 
of interest. The bane of shallow seismic 
work is the fact that low-frequency 
ground roll predominates in the time win- 
dow of the desired reflections. Its 
effect is to swamp the amplifiers, and 
their effective dynamic range, before 
wanted signals can be recorded at reason- 
able signal-to-noise levels. Mere stack- 
ing of repeat shots, or filtering after 
the fact at the computer center, will not 
help. In the first case, the ground roll 
will be coherent for repeat shots; and 
weak signals embedded in large groundroll 
excursions usually cannot be effectively 
extracted from the data. 



16 



SURFACE DISTANCE, ft 



0.1 — 



ivvmvJ 







'"^^ 






Coal reflector 
position 






Jilt - 




X-Ocx;-4-^ 



VvXa: 



C-C C <. C <'<.>:«; i:i.v IV inn^^^rT^^^PP^>VVkWfc.w 



— .2 






FIGURE 10, = Seismic profile of a coal measure, Colorado. 



SURFACE DISTANCE, ft 



17 





o 


o 


o 
o 


o 


o 


CM 


lO 


o> 


T— 


T- 



- 



BUREAU OF 
MINES 



0.1- 



.3- 




iiirMin;'" 



"'••Mini' 

,,' '•►►►►►►►►HI 
>»»»»»»,. >>>>> 




0.1 



.2 



r- -3 



EBENSBERG PH 
FIELD DRTR 
SHOT JUNE 4 1976 
I 8 TR 1 2 MS 
PHONE SPRCINC 50 
SOURCE OFFSFT IDO 
CHRRGE SIZE 1 3 
PROCESSINC ORTR 
DRTUM ♦ZOOO vCSOOO 

boh; notch filter 

GROUND ROLL MUTE 
INV NOISE FILTERS 
VELOCITY FUNCTION 
000-5500 2G0-9250 
000-6250 294-i02S0 
0"0-B750 310-11250 
120-6850 392-12250 
180--250 999-10000 
230-8250 

FILTER 80-240 HGC 
SEM MODELING 

RUTOSTHCK 
FILTER 50-lSO RGC 
DIRECTION 

STR 2 NE TO 37 



.6- ;:::;; 



FIGURE 11. - Seismic profile of a coal measure, Pennsylvania, 



In addition to ground roll, there may 
be other peculiarities of the site, such 
as a high-velocity layer near the sur- 
face, or spurious reflections from a 
nearby cliff face or stream cut, or vi- 
brations from nearby machinery, or traf- 
fic (if in an urban area or near an 
active mining site). Actually, the 
sometimes reported "seismically blind" 
area seldom exists. What is missing is a 
full understanding of the seismic nuances 
and the tools at hand. Bureau experience 
has shown the importance of having a 
repertoire of equipment and potential 
approaches, and the willingness to take 
the time to test and characterize the 
area. The field crew, and their knowl- 
edgeable geophysicist, should not be 
pressured into forging ahead into the 
production mode before these tests are 
made and the results field analyzed. A 
week, or even two, if on a large pros- 
pect, will be worth the investment. 

Some initial decisions have to be made. 
What is the desired time window of the 
zone of interest? If it's from, say, 300 
to 500 ft, then it is important to test 
for the optimum field configuration that 
will give the best signal-to-noise ratio 
in that time window. Although each oper- 
ator's technique may vary. Bureau re- 
searchers have found that a good first 
step is to lay out a close order spread, 
with the geophones clustered (if subar- 
rays are used) at 1-m spacing. Then one 
should produce energy sources at 1 m from 
the nearest phone, then at one spread 
length, then two spread lengths, then N 
spread lengths away, until the ground 
roll is obviously coming in far beyond 



the time zone of interest. This layout 
is shown in figure 12, and a set of 
resulting seismograms in figure 13. The 
layout figure is typical only, and should 
be varied in accordance with personal 
experience and preference. 

The seismogram set of figure 13 shows 
many things. First of all, groundroll 
frequency (in this case 27 Hz) is now 
known and can be used for spatial filter- 
ing judgments of either the subarrays (if 
used), the signal source pattern, or 
both. Farr (5^) discusses this procedure. 
The presence of the reflector at 175 msec 
indicates that a shot to the first geo- 
phone offset-distance of 200 ft will 
assure its continuing reflection at rea- 
sonable signal-to-noise levels. The 
spurious reflection, band A, coming in 
around 150 msec, indicates that noise 
vibrations are coming into the spread 
from the nearby industrial area. Its 
predominant 125-Hz frequency indicates 
that it can be subdued, either by spatial 
filtering of the geophones in the sub- 
array, or perhaps by pattern shooting 
approximately perpendicular to the line, 
or ideally in the direction of the noise 
source. If this type of source is fixed, 
its location can be determined from the 
test spread with a little geometry. If 
it is a moving source, then it becomes 
necessary sometimes to delay firing until 
such noises subdue. 

During the generation of the test 
spread, one should perform selective gain 
setting changes. From this, a typical 
range of gain settings can be determined 
for most of the ensuing survey. 



0.5m^ \^ ^ 2m|^-4m-H 



Geophone 
spread 



Shot points 
Continue to approximately 3 times spread length- 



• Individual geophones 

^ Shot points into geophone spread 

FIGURE 12. = A noise spread layout. 



19 



From the figure 13 example, which was 
in a location declared "seismically 
blind," some reasonable first approxima- 
tions of field layout parameters re- 
sulted. Had this not occurred, it would 
have been necessary to reshoot each shot 
point with a best judgment weighted pat- 
tern configuration, using Farr's work (_5) 
and ground-roll frequencies as a guide. 
If a pattern shooting had additionally 
been done, the resulting section would 
aid in defining an optimum field proce- 
dure. In retrospect, the direction of 



the spurious reflection at 150 msec could 
have been determined by a concentric cir- 
cle geophone pattern (I) . This layout 
would have established the direction of 
the spurious reflection and further 
influenced the decision towards an opti- 
mum field procedure. From the first 
breaks. Bureau researchers could deter- 
mine much about, the velocity layers near 
the surface. This helped in establishing 
judgments on whether or not sharp veloc- 
ity discontinuities existed and an ap- 
proximation of weathering depths. 



Band "A" o'- tHi"^**' 



1 50 msec 

First 
reflector,^ 



X: 

1 1 1 1 > il I JiitMH'' 



KfcW__J.^tttKU^' 



-i^M.rt'rttliiK'i 



MiJMIlMlitih 



^iHHM 



nlHiii 



175 msec .L^^^^^^^^^,^I»^H^ 




400 350 300 250 200 150 

DISTANCE, ft 



00 50 



S <- -, 5 



UNITED STATES 



BUREAU 0F MINES S 



SUSANVILLE 
CAL IF0RNI A 



i N015E PROFILE i 

s ft 

i FIELD DATA i 



SH0T BY U S B M MAY 1977 
U0 8 CHANNEL DINeSEIS SURFACE 
TRACE INT 6 25' SP 50' 1/4 nS 51 

14 HZ GE0PH0NES 



esssssssssssssssssssssssssssssse 



FIGURE 13. - A noise spread profile. 



20 



At this point, one will want to decide 
what further procedures to try to improve 
signal-to-noise conditions. This always 
must remain No. 1 in one's mind. It is 
also important that reflections must be 
seen in the data, even if only cursory at 
this point, before signal enhancement 
procedures, such as Common Depth Point 



(CDP) or signal averaging (stacking), 
will aid the quest. If reflections can 
be seen through the procedures outlined 
above, one can feel confident that CDP 
and stacking will help. If they can't be 
seen, then one should reassess offset, 
geophone, or shop pattern layout dimen- 
sions, or perhaps even the signal source. 



3-D SHOOTING TECHNIQUES 



Many examples of the effectiveness of 
3-D techniques are to be found in oil 
exploration literature. Some work has 
been performed for mining application in 
Europe, and increasingly in the United 
States. Where it has been used in the 
United States for mining, the results 
have been exceptionally good. A few 
examples of this work are presented in 
this section; some are by seismic indus- 
try companies, and some by Bureau re- 
search personnel. 

The case for considering 3-D techniques 
revolves around their inherent ability to 
provide thorough coverage of a site in 
all dimensions. High-resolution tech- 
niques may often mean just high-density 
collection of noise. Recovering shallow 
information requires multiple data 
schemes. The simplest, and most often 
used, has close-spaced traces, 20 to 
25 ft apart, with single 40-, 60- or 100- 
cycle geophones or hydrophones in water- 
filled 10-ft holes, and a small charge 
detonated in a 1- to 10-ft drilled hole. 
This scheme rarely achieves the objec- 
tives, because ground roll is broad band, 
and exploding concentrated charges within 
the poor elastic shallow environment ex- 
ceeds the elastic limits and is poor in 
the high-frequency range. A successful 
procedure mast select an operating fre- 
quency band in which (1) source and 
receiver arrays can attenuate ground roll 
within the physical limit rendered by 
depth or by dip reflection objectives, 
and (2) a high-velocity explosion or de- 
vice is used in conjunction with low-unit 
loading that will not exceed the elastic 
limits of the surface environment. 

Further, to fulfill the array- 
attenuation procedures, source and re- 
ceiver patterns will need to be minimally 
AO to 10 ft or longer. The physical 



requirements of ground-roll suppression 
are difficult, or perhaps impossible, 
with conventional profiling procedures, 
if the objective is shallow or dipping. 
The depth, and to some degree the veloc- 
ity, limits the maximum distances from 
source to receiver. A good rule of thumb 
is that the longest offset be limited to 
the depth of the shallowest reflecting 
objective. Therefore, to map a 500-ft 
event with stations spaced 100 ft apart 
allows only 10 to 12 traces of informa- 
tion, severely restricting the CDP fold. 
The 3-D procedure overcomes these re- 
strictions. Using the 3-D matrices of 
source and receiver stations permits al- 
most unlimited array lengths. Also, the 
pattern can be oriented along strike and 
data collected in the dip direction. 
With 3-D procedures, 12, 24, 96, or more 
CDP folds can be recorded. 

In seismic profiling, the geophones and 
signal sources are arranged as shown in 
figure 14. The resultant seismic profile 



Energy source 
n" , X 7 Geophones 



AV 



lmUi 



(To coal or shollow 



ii 



FIGURE 14. - Depth of shallowest reflector versus 
maximum distance to furthest geophone. 



21 



will be as shown in figures 4, 5, 6, or 
10. These profiles, though valuable, 
represent only a single cross section of 
the coal geology. 

A 3-D survey, on the other hand, pro- 
vides a multitude of profiles, and in 
addition, a series of time-depth slices 
that enables one to examine the ore 
deposit geology layer by layer. Layout 
configurations are limited only by the 
geophysicist's imagination and surface 
constraints, such as hills, bodies of 
water, or cultural development. All of 
these can be effectively surveyed with 
3-D procedures. 

Where conditions permit, a rectangular 
layout, such as shown in figure 15, is 
cost effective. The active spreads are 
progressively advanced over the grid. 
Energy inputs are generated on an 



overlapping grid that starts at a dis- 
tance "d" from the first active spread 
line. The distance "d" is chosen from 
judgments of previous surface noise tests 
and should be a multiple of the longitu- 
dinal grid spacing. Each grid position 
is energized by the energy source, 
whether explosives or surface impactors. 
The consequent subsurface coverage is 
prolific. This is shown in figure 16, 
wherein a controlled chamber explosive 
source was used for rapid coverage. Al- 
though simple CDP coverage from 3-fold to 
12-fold was immediately obtained by com- 
puter gathering of the data, this was 
increased from 27-fold to 72-fold with no 
additional field effort. Figure 17 shows 
one line profile taken from the data. 
Additional profiles may be along any 
other horizontal or longitudinal line, or 
in any direction. The potential of this 
thorough coverage should be obvious. 



Detai 
"A" 



¥: )i ¥r ¥r ¥:¥: ¥:¥r ¥: ^^ )i^ F i r s t s h 1 poi 1 1 i n e 

(••••••••••••T~) '5' active spreod ^ 

-^^-p-gB ■■■■■■■ ■ ^ 2d active spread '^ Lines may be single 
▲ AAAAAAAAAAA ordual(shown) 

AAAAAAAAAAAA l^^ CCtlve SpreOd I 



Hi 



NOTE d IS *ti^ oftsef distonce from 
energy source line to nearest 
And shotpoint geophone line 

spread progression 



Optimi7ed 
choice from 
noise tests 
should be y 
factor of ' d' 



A s-ngle geophone 
a subarray (shown) 
may be used 



No of active in-lne geophones 




FIGURE 15. = Plan view of a rectangular 3=D survey layout. 



22 



tZ'"?ld 3-D CaVERAGE 

SUB-SUF^FACE 

27 333333333 

48 366666663 

63 369999963 

72 3 6 9 12 12 12 9 6 3 

72 3 6 9 12 12 12 9 6 3 

72 3 6 9 12 12 12 9 6 3 

72 3 6 9 12 12 12 9 6 3 

72 3 6 9 12 12 12 9 6 3 pian vjew 



of seismic 
grid 

72 3 6 9 12 12 12 9 6 3 

1-72 — 3- -6 — 9--12— 12--12--9--6--3 -J Profile 

A-A' 

72 3 6 9 12 12 12 9 6 3 



72 3 6 9 12 12 12 9 6 3 



72 3 6 9 12 12 12 9 6 3 

72 3 S 9 12 12 12 9 6 3 

72 3 6 9 12 12 12 9 6 3 

72 3 6 9 12 12 12 9 6 3 

72 3 ^ 9 12 12 12 -^ 6 3 

72 3 6 9 12 12 12 9 6 3 

72 3 6 9 12 12 12 9 6 3 

72 3 6 9 12 12 12 9 6 3 

■72 3 6 9 12 12 12 9 6 3 

S3 369999963 

4? 366666663 

_^ 333333333 

FIGURE 16. - 3~D grid showing immediate fold 
and optimum total fold. 

A method of present:ation for these data 
has been evolved by several seismic data 
processors, wherein time-depth slices are 
taken through the zone of interest, as 
shown in figure 18 for a coal deposit. 
Here the top of the coal is black, and 
coal itself is white. The interpretation 



SURFACE DISTANCE, ft 

o 
o 

(N O 



AQUIFER 



Hanna#1 coal 
Hanna#2 coal 02 i 



FORT UNION 



CRETACEOUS 
SHALE 




08 § -fff^|i54^i|«^^l|if 



09 

10 I 

11 ^ 

12 ^ 

13 ^ 

14 - 



" T^^^^W^?<ftt^ = 09 



;^. ' : ' r' ' r:{^r% > ' > H S §11 



4f^ 



r^ >» > >i^,>>M(>» 



14 



JURASSIC 






15 u»,;„;vm>>v,»>»*;^» ^ 15 

16 -:;::::; ;;;:::!;[:;:::::: ne 




.»'*>|>>,»" ' K»l»l>t^>> 




17 - .,'»'F*'(»».vH^>»»^»»f>» - 17 




»»*M',»t'>»Mt»»'»M»> - 




18 :f^^ 18 




19 - :.|f ePTinuM ftr : 19 

||llAPPAr STACKING l|l| 

20 - illlllllllHIIIIIIllllli ^20 




A -—profile A' 



FIGURE 17. 



Typical time=depth slice through 
center of project. 



shows the undulations and dip of the 
seam, and a possible fault cutting across 
it. These features would have been dif- 
ficult to ascertain from the profiles 
alone, but become quite evident on the 
time-slice presentation. Coal depth is 
about 270 ft. 



23 




FIGURE 18 - Typical time slice of 3=D seismic. 



Figures 19 and 20 provide another exam- 
ple of this procedure using an 8-channel 
system for recording and a surface impac- 
tor signal source for a coal deposit. 
Coal depth is about 180 ft. Of signifi- 
cance in this example is the high qual- 
ity of seismic data throughout the sec- 
tion from above the coal to over 2,000 ft 
deep. Although no abrupt geologic fea- 
tures occur in this coal section, its 
dipping subtleties can be seen. Had 
there been any abrupt change, they would 
have shown up readily. 

In another area in southeast Kentucky, 
a modified 3-D survey was conducted. 
Because of terrain complexities, the 
seismic procedure was essentially "fan 
shooting," but it provided a 3-D coverage 
over much of the area. The coverage 



shown is in a hill-valley setting shot in 
midwinter, which was impassable except 
along bulldozed roadway. Figure 21 shows 
the coverage obtained in one section of 
the prospect. Note here that access 
could only be obtained along the roadway, 
as identified by the shotpoint numbers, 
yet coverage of the site included all of 
the areas shown in the reflection point 
map of figure 21. By taking selected 
data gathers through this covered region, 
a series of profiles was obtained along 
the road and in many of the inaccessible 
areas. From these, the top of the coal 
seam could be mapped. 

Although the 3-D procedure is more 
costly than profile techniques alone, it 
provides excellent total coverage data 
that cannot be obtained by any other 



24 



GEOPHONE POSITIONS 




WW: 




vl^.iiivvi 


^f§ 


•j'j."^*^ 


>tili;;:; 


^;;;;;>> 


~;i;;;; 


""::::: 


"^■li-Hii 


,>.•.') J J.;?/ 


ii!::!'' 








U».jJ--UJ 


iltV!!*)J 


lii»"t^ 


"""v'/r 


v„v,,vv. 


►..►..«. 


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.►.►v.... 


«^».,. 


fi!:[;::: 


ii,''rr"j 


l[^'^^ 


rrrffrff: 


i::::::ff 



FIGURE 19. - Seismic 3-D profiles using an 8- 
channel recording system. 



M 




TOP OF COAL REFLECTION SECTION, at 167 msec 








MARKER BED BELOW COAL, at 195 msec 

FIGURE 20. - Timeslicefrom3=Dsurvey. 

means. It is recommended for difficult 
areas where profiling alone seems to 
fail, or where thorough coverage is want- 
ed to accurately define ore deposit con- 
ditions. The trade-off is an investment 
in excellent data versus merely suffi- 
cient data to outline a potential problem 
area. 



CONCENTRIC CIRCLE TECHNIQUES 



At times, there are problem areas that 
require special attention. This may be 
because of high-dip conditions, a near- 
surface high-velocity layer, an adjacent 
large fault, steep hills, valleys, cliff 
faces, or cultural noise. A powerful 
procedure for tackling these areas is a 
concentric-circle technique developed by 
several oil companies, and documented by 
Brasel (_1_). The power of the concentric 
circle technique is its ability to act as 
a directional antenna, from which one may 
readily determine the direction of spuri- 
ous noises, and hence, separate them from 



the wanted reflections. The spreads are 
relatively easy to lay out and can run as 
a single continuous trace for line- 
profiling across complex terrain. Conse- 
quently, it is effective in areas where 
neither normal spreads nor 3-D techniques 
can be used. This is shown hypotheti- 
cally in figure 22. Here it is presumed 
that most of the area is inaccessible for 
3-D or continuous profiling. If the con- 
centric circle spreads are laid out 
wherever possible, a reasonable continu- 
ous profile, having good-quality data, 
can be obtained. 



25 






jj , o . . . c . « . <, 



30 



It 



•Subsurface coverage 
(reflection points) 

-Subsurface Interpretation-grid 
locations and reference numbers 



Geophone and 

shotpoint location 

and reference numbers' 



Geophone and shotpoint locations 



Note: Subsurface interpretation— gride locations. 
FIGURE 21 - Seismic 3=D coverage in hilly country, Kentucky, 




Typical diameter 24ft (lOm) 

note: L is typically I to 3 diameters 
as dictated by site condition 

FIGURE 22. = Concentric circle traverse layout in difficult terrain. 



Concentric circle spread layouts 



26 



Layout dimensions are site dependent, 
but some layouts that Bureau researchers 
have had experience with are shown In 
figure 23. In this example coal depths 
were approximately 200 ft (60 m) . Each 



geophone position is a separate trace; 
hence, CDP stacking could be performed. 
Quality reflections are obtained at each 
position that could be readily correlated 
from spread position to spread position. 



COMMON DEPTH POINT (CDP) 



This is a powerful tool for enhancing 
signal-to-noise ratios. Its use should 
not be considered as a crutch, however, 
to bail out dissatisfaction from unwel- 
come results from the Initial test pro- 
cedure of the previous section. CDP pro- 
cedures will Improve the final quality of 
the data, but will not find reflections 
that are deeply embedded in noise. This, 
in part, is due to the Inherent reality 
of where the CDP Information comes from, 
and the coii5)uter-procedure presumed 



location. The two locations are close, 
but not exact. To explain further: 

CDP procedures, always a conputer 
process, are based upon a powerful 
data-gathering technique (8-^) , wherein 
reflections are obtained from the same 
location at different shotpolnt versus 
geophone positions. This is shown di- 
agramatlcally in figure 24. The figure 
shows that multiple reflections from a 
single point can be obtained from 






Test 
configuration 




J_._._ 



34ft(IOm)-»| 



Traverse 
configuration 



2 diam 



V 

Test shot points each at I diam. separation 





• 


Geophones 
Shotpoints 




B 


V 


C 


D 
•- 


J 







A,B,C,D, ondE ore geophone clusters as 
shown in the test configuration 



FIGURE 23. - Typical concentric circle traverse layout for difficult areas. 



27 



600pct multiplicity 
12 stacking channels-split array 

Subsurface coverage 



Surface coverage 



V^ 



Energy travel paths 



Subsurface coverage 
from shot points 123456 




SP 

No. 


Depth points | 


I 


2 


3 


4 


5 


6 


7 


8 


9 


10 


II 




1 


II 


12 






















2 


9 


10 


1 1 


12 


















3 


7 


8 


9 


10 


II 


12 














4 


6 


- 


7 


8 


9 


10 


II 


12 










5 


4 


5 


6 


- 


7 


8 


9 


10 


II 


12 






6 


2 


3 


4 


5 


6 


- 


7 


8 


9 


10 


1 1 


12 


7 




1 


2 


3 


4 


5 


6 


_ 


7 


8 


9 


10 


8 








1 


2 


3 


4 


5 


6 


- 


7 


8 


9 












1 


2 


3 


4 


5 


6 




10 
















1 


2 


3 


4 


5 



"—"indicates no suburface 
coverage obtained 

¥r = Shot points 
SP No. = Shot point number 



FIGURE 24, - Stacking configurations. 



geophone-shotpoint combinations when a 
line is shot in a roll-along mode. When 
conputer summed, these add to signal- 
to-noise enhancement. The exact posi- 
tion, however, is a statistical variable, 
dependent upon travel path parameters, 
such as varying frequencies, angle of 
reflection, scattering, anistropy, and 
conplex signal propagation considera- 
tions. Hence, it can smear frequencies 
somewhat. None of this discounts its in- 
herent value, however, if used knowingly. 



In shallow seismic reflection work, one 
should progressively test-select gathers 
from a good reflector to see whether the 
final data are actually enhanced or 
degraded before arbitrarily gathering the 
maximum possible combinations. 

Bureau of Mines work has shown that a 
6- to 12-fold GDP stack is usually suffi- 
cient to enhance most shallow reflection 
data. 



CONCLUSIONS 



Shallow seismic techniques are useful 
for many hard rock and sedimentary mine 
planning purposes. Much useful data can 
be obtained from these procedures. When 
they are supplemented with selective 
drilling, an enhanced picture of subsur- 
face geologic conditions can be obtained. 
The seismic data fill in the picture 
between drill holes and provide a clearer 
understanding of the geology than can be 
obtained by drilling alone. 

Although many of the procedures are 
still in development, enough has been 



learned in the past 6 years to show that 
the technology is viable for mining. 
Many mining companies are fielding their 
own crews or using the services of con- 
tract seismic companies. Seismic ser- 
vice companies are increasingly offer- 
ing their new expertise to the industry. 
The essence of the use remains in 
conducting sufficient a priori seismic 
tests to obtain the best signal- 
to-noise window for the depth zone of 
interest. 



28 



The linear seismic line is useful for 
many applications. For total comprehen- 
sive coverage, the 3-D approach is a 
must. Commercial costs for linear work 
will be approximately $100 per shot 
point, while 3-D coverage will be approx- 
imately $40 per grid position. Crew size 



for this work ranges from 6 to 12 per- 
sons, of which one is at an engineering 
level. From the point of view of the 
potential value to a mining operation for 
both safety and extraction economics, 
these costs are nominal. 



REFERENCES 



1. Brasel, S. D. Circular Arrays 
Applied to Conventional and 3-D Seismic 
Surveys. Seismic Res. Internat. , July 
1978, p. 20. 

2. Dix, C. H. Seismic Prospecting for 
Oil. Harper and Brothers, New York, 
1957, p. 414. 

3. Dobecki, T. L. , and S. D. Brasel. 
Aerial (3-D) Seismic Applications to Coal 
Seam Characterization and In-Situ Gasi- 
fication Projects. Preprint from SEC 
48th Internat. Symp., Dallas, Tex., Geo- 
physics, Society of Exploration Geophys- 
icists, Tulsa, Okla. , September 1978, 
p. 15. 

4. Embree, P., and others. Wire-Brank 
Velocity Filtering — the Pie-Slice Pro- 
cess. Geophysics, Society of Explor- 
ation Geophysicists. V. 28, 1963, 
pp. 948-974. 

5. Farr, J. Seismic Profiling for 
Coal Mine Planning. Geophysics, Society 
of Exploration Geophysicists, Tulsa, 
Okla., V. 44, No. 3, 1979, p. 324. 



Concepts of CDP and Digital Processing. 
Geophysics, v. 32, 1967, pp. 207-224. 

9. Mayne, W. H. Common Reflec- 
tion Point Horizontal Data Stacking 
Techniques. Geophysics, v. 77, 1961, 
pp. 927-938. 

10. Pakiser, L. C, and R. E. Warrick. 
A Preliminary Evaluation of the Shallow 
Reflection Seismograph. Geophysics, 
V. 21, 1956, pp. 388-405. 

11. Sauit, C. H., and others. The 
Moveout Filter- Geophysics, v. 23, 1958, 
pp. 1-25. 

12. Schneider, W. A. Developments in 
Seismic Data Processing and Analysis. 
Geophysics, v. 36, 1970, pp. 1043-1073. 

13. Sheriff, R. E. Encyclopedic Dic- 
tionary of Exploration Geophysics. So- 
ciety of Exploration Geophysicists, Tul- 
sa, Okla. , 1974, 200 pp. 

14. White, J. E. Seismic Waves. 
McGraw-Hill, New York, 1965, p. 220. 



6. Hasbrouck, W. P. Instrumentation 
for Coal Seismic System. Geophysics, 
Society of Exploration Geophysi- 
cists, Tulsa, Okla., v. 44, No. 3, 1979, 
p. 377. 

7. Lepper, C. M. Guidelines for Se- 
lecting Seismic Detectors for High Reso- 
lution Applications. BuMines RI 8599, 
1982, 72 pp. 



15. 



Patterns, 
pp. 26-43. 



Transient Behavior of 
Geophysics, v. 23, 1967, 



16. Ziolkowski, A., and W. E. Lerwill. 
A Simple Approach to High Resolution 
Seismic Profiling for Coal. J. European 
Assoc. Exploration Geophys. , London, 
England, May 1977, 26 pp. 



8. Marr, J. D. , and E. F. Zagst. Ex- 
ploration Horizons From New Seismic 



29 



APPLICATION OF THE ELECTRICAL RESISTIVITY METHOD TO MINING PROBLEMS 
By Richard G, Burdickl 



ABSTRACT 



Electrical resistivity methods have 
been used for a variety of mining appli- 
cations in the past, and current re- 
search is being directed towards fur- 
ther applications to premining hazard 



detection and monitoring solution mining 
areas. Some of the methods are dis- 
cussed with examples of past mining 
applications. 



INTRODUCTION 



The use of the electrical resistivity 
methods in mining situations is finding 
expanding application. They may be used 
to locate geologic faults, measure the 
degree of fracturing within in-place 
rock, locate abandoned mine workings or 
tunnels, and define the limits of solu- 
tion fronts used in solution or in situ 
mining. The method is not a panacea for 
all mining problems, but, when used on a 
case-by-case basis, it is a versatile 
tool for a wide range of mining problems. 
It has the advantage that it may be used 
from the surface or underground as well 
as in drill holes. 

The two resistivity methods most com- 
monly used by the Bureau for mining 
applications are the Wenner method 
(appendix A) and the pole-dipole method 
(appendix B) (j_, 4_, 8^, j^) . In general, 
the Wenner method is used to investigate 
larger volumes of earth materials where 
its inherent averaging effect can be used 
to advantage. The pole-dipole method is 
used where more discrete samplings of the 
earth are required, particularly at 
greater depths. This paper describes 
applications of resistivity methods to 
the solution of various mining problems 
by the Bureau. 

Discussion of Resistivity Principles 

As with other geophysical methods, re- 
sistivity does not detect the target- 
of-interest directly, but rather measures 

^Engineering technician, Denver Re- 
search Center, Bureau of Mines, Denver, 
Colo. 



the electrical properties in an area, and 
from these an inference may be drawn as 
to the presence or location of the 
target. 

Even the most competent rock contains 
small voids or microfractures that con- 
tain water and various dissolved salts 
and gases. These solutions allow the 
passage of electrical current and form 
the basis for the resistivity measure- 
ments. A dense, conqjetent rock with few 
microfractures will contain relatively 
less moisture than a more porous rock and 
will normally exhibit a much higher re- 
sistance to the passage of an electrical 
current than will the more porous rock. 
Thus, the resistivity methods do not mea- 
sure solution fronts, abandoned mine 
workings, etc., directly, but simply mea- 
sure the changes in the volume resistiv- 
ity caused by their presence. 

The equipment for making these measure- 
ments can be as simple as using a battery 
for injecting current into the ground and 
a sensitive voltmeter to measure the re- 
sulting potential, or as complicated as 
the Bureau's prototype Automatic Resis- 
tivity System, which closely controls the 
current injected and has an elaborate 
measurement system that accurately mea- 
sures to 10 mv (0.00001 volt) and records 
all data, including measurement loca- 
tions, on magnetic tape for later com- 
puter analysis. The intermediate range 
equipment includes many commercially 
available units, as well as somewhat more 
sensitive equipment designed as prototype 
devices (5). 



30 



The apparent resistivity of a volume of 
earth using the Wenner method is calcu- 
lated by the formula: 



AV 



app 



where 



,pp = apparent resistivity, in 
ohm-meters. 

I = current injected into the 
ground, in amperes. 

AV = the voltage potential 

between two electrodes 
resulting from the above 
current. 

k = a geometry factor of 2ira. 

a = electrode spacing, in 
meters. 

From this equation it may be seen that 
when the electrode spacing is kept con- 
stant there is a direct relationship 
between apparent resistivity and the mea- 
sured AV/I. If the earth moisture 
changes there will be a corresponding 
change in the measured AV/I ratio. 

Use of Resistivity To Predict 
Caveability of Ore Bodies 

As a fracture zone within an ore body 
below the water table would be expected 
to contain relatively more moisture than 
the surrounding ore, it would also be 
expected to show a lower resistivity than 
the other ore. Based upon this premise, 
studies were conducted at two mines using 
block-caving extraction methods. Four or 
five separate areas were tested and rela- 
tive caveability values (best, medium, 
and poor) were assigned based upon the 
relative apparent resistivities of the 
areas. In general, these predicted rela- 
tive caveability values were borne out 
when the areas were mined at a later date 
(Source: Unreported Bureau investiga- 
tions during 1966-68). 



Use of Resistivity To Define the Zone 
of Fracturing Around a Tunnel 

Based upon the general logic that is 
the greater the degree of fracturing the 
lower the apparent resistivity, a brief 
study at one hardrock mine seemed to show 
the zone of fracturing occurring around a 
tunnel blasted through the rock. In this 
case, the Wenner Array sounding method 
was used as described in appendix A. The 
electrode spacings were expanded in small 
increments in order to detect the edge of 
the fracture zone. From this test, the 
fracture zone appeared to extend to 5 ft 
out from the walls of the tunnel (Source: 
Unreported Bureau experiments). 

Use of Resistivity To Locate 
Geologic Faults 

The Bureau has used the resistivity 
traverse method on several occasions to 
determine the surface location of faults 
in mining areas. The principles used in 
this type investigation are shown in fig- 
ure 1. The fault may be detected either 
if it serves as a drain for surface mois- 
ture resulting in a somewhat drier zone 
around the fault, or if it tends to con- 
tain more moisture than the surrounding 
area. If the fault has shown enough ver- 
tical displacement to bring dissimilar 
lithologies near enough to the surface to 
be within the measurement zone, a third 
possibility for detection exists. All 
three phenomena have been observed in 
the past. The method has been used 
in two uranium fields in Wyoming and at a 
hardrock mine in Colorado with good 
results (9^). 

Use of Resistivity for Detection 

of Abandoned Mines in Proximity 

to Current Mining Activities 

Both the Wenner and pole-dipole methods 
have been used for the detection of aban- 
doned mines. However, the pole-dipole 
method shows the better resolution for 
for this purpose, because the Wenner 



31 



Moisture differences 



Lithologic differences 



/Ground surface -. 
If- ^ 



II 
II 
I, ^^ Fault rone 

/A 

// 
// 




GEOLOGIC PROFILES 



drier 

Fault J \ than surrounding 



rock 



wetter 



Curve inflection caused 
by dissimilar lithologies 



TRAVERSE POSITION 



TRAVERSE POSITION 



FIGURE L - Use of resistivity to detect faults. 



measurement volume is much larger for a 
given depth of investigation than the 
pole-dipole and therefore is not as sen- 
sitive to detection of a void. This is 
shown by figures A-2 and B-1. The prem- 
ise made for this type investigation is 
that the old workings, whether air- or 
water-filled, will show a different 
apparent resistivity than the surrounding 
earth material. Figure 2 shows the use 
of the pole-dipole method to detect a 
void. The lines A, B, and C would be run 
at the same time with the automated sys- 
tem. From the points where the void was 
detected on A, B, and C, the void may be 
located in a manner similar to that shown 
on the lower right figure. Either the 
pole-dipole or the Wenner methods have 
been used in Florida, Illinois, Kentucky, 
Colorado, Korea, and Israel for locating 
various underground voids, usually with a 
high degree of success. Further work is 
being done in this field to try to in- 
crease the ability of the method to 
detect smaller tunnels at greater depths 
(7_). The Bureau's automated system is 
still in a testing and modification 
stage. It is anticipated, however, that 
the method will be fully operational in 
about a year and, after that time, a 
similar system could be constructed by 
interested parties or by the original 
contractor for such parties. 



Current electrode 




00 

Current electrode 



LINE A DATA CURVE 



Line A data 













Line B da 






Line C data curve 


1^ 




C B 




FIGURE 2. - Void detection with pole-dipole method. 

Use of Resistivity To Predict Roof Falls 

The Bureau has been conducting research 
for several years, both in-house and 
under contract (2^), to devise an auto- 
mated roof fall warning device using 
resistivity principles. The assumption 
made in this case is that the delamina- 
tion or fracturing cracks preceding a 
fall will cause a change in volume appar- 
ent resistivity, which can be interpreted 
as an indication of failure (fig. 3). 



32 




\ /Al Measurement field //,'', 

.; 'v \\\ W ) 




, 



^^^^^^^ Ground surface 

^ '^ p 




ll /"^ ■* Solution front 





Potential electrodes 




^al of solution front 



FIGURE 3. - Use of resistivity to predict roof falls. 



FIGURE 4- • Use of resistivity to monitor solu- 
tion fronts. 



This figure shows how cracks caused by 
de lamination or other fracturing can 
affect the resistivity measurements. The 
plot of resistivity versus time shows the 
normal scatter of data points preceding 
failure and the much greater change in 
values as cracking starts and progresses. 

Use of Resistivity To Monitor Solution 
Fronts in In Situ Extraction of Ores 

The Bureau of Mines' Twin Cities Re- 
search Center has recently con5)leted a 
contracted research effort in which the 
solutions or lixiviants used for uranium 
mining were monitored as they approached 



the edge of the mine area. This method 
can be used to monitor the progress of 
solution mining as well as to detect the 
unwanted migration of the solution beyond 
the extraction zone. The method relies 
on the fact that the conductivity of the 
solution is quite dissimilar from that of 
the natural moisture in the rock, and 
when it flows into the measurement field, 
a change in apparent resistivity will 
occur. The method has the advantage that 
the solution movement can be detected in 
the zone between drill holes rather than 
just at the drill hole, as had been the 
case in the past. Figure 4 shows in gen- 
eral how the technique operates (3). 



CONCLUSIONS 



The described uses of various resistiv- 
ity methods in a number of mining envi- 
ronments show their versatility for solv- 
ing a variety of problems. One of the 
method's strong points is its adaptabil- 
ity for use on the surface or in an 



underground environment. Another is the 
simplicity of making measurements and 
interpreting the resulting data. In many 
cases, this means that an interpretation 
of the data may be made while at the test 
site. 



REFERENCES2 



33 



1. Dobrin, M. D, Introduction to Geo- 
physical Prospecting. McGraw-Hill Book 
Co., New York, 1952. 

2. Gibbons, M. , A. J. Farstad, and 
R. F. Kehrman. Resistivity Roof Fall 
Warning System in the White Pine Mine. 
BuMines Open File Rept. 53-80, 1979, 
75 pp., contract H0272037, Westinghouse 
Electric Corp.; available from National 
Technical Information Service, Spring- 
field, Va., PB 80-186547. 

3. Kehrman, R. F. Detection or Lixiv- 
iant Excursions with Geophysical Resis- 
tance Measurements During In Situ Uranium 
Leaching. BuMines Open File Rept. 5-81, 
1979, 156 pp.; contract J0188080, West- 
inghouse Electric Corp.; available from 
National Technical Information Service, 
Springfield, Va. , PB 81-171324. 

4. Koefoed, 0. Geosounding Princi- 
ples. Elsevier Pub. Corp. , North Hol- 
land, Netherlands, v. 1, 1976, 276 pp. 

5. Lepper, C. M. and J. H. Scott. An 
Improved Electrical Resistivity Field 

2 Items 2, 3, and 7 are available for 
reference at the Denver Research Center, 
Bureau of Mines, Denver, Colo. 



System for Shallow Earth Measurements. 
BuMines RI 7942, 1974, 20 pp. 

6. Orellano, E., and H. M. Mooney. 
Master Tables and Curves for Vertical 
Electrical Sounding Over Layered Struc- 
tures. Interciencia, Madrid, Spain, 
1966, 234 pp. 

7. Peters, W. R. Detection of Coal 
Mine Workings Using High Resolution Earth 
Resistivity Techniques. BuMines Open 
File Rept. 55-81, 1980, 70 pp.; contract 
H0292030, Southwest Res. Inst.; available 
from National Technical Information Ser- 
vice, Springfield, Va. , PB 81-215378. 

8. Sharma, P. V. Geophysical Methods 
in Geology. Elsevier Pub. Corp., North 
Holland, Netherlands, 1976, 427 pp. 

9. Stahl, R. L. Detection and Delin- 
eation of Faults by Surface Resistivity 
Measurements, Gas Hills Region, Fremont 
and Natrona Counties, Wyo. BuMines 
RI 7824, 1973, 28 pp. 

10. Van Nostrand, R. G. , and K. L. 
Cook. Interpretation of Resistivity 
Data. U.S. Geol. Survey Prof. Paper 499, 
1966, 310 pp. 



34 



APPENDIX A.— DESCRIPTION AND INTERPRETATION OF THE WENNER METHOD 



The Wenner method uses a configuration 
of four equally spaced electrodes, as 
shown in figure A-1. The current is in- 
jected between electrodes C, and C2, and 
the potential difference is measured be- 
tween electrodes P, and P2. Figure A-2 
shows the approximate current and poten- 
tial fields for a Wenner electrode con- 
figuration. It is generally assumed that 
the potential being measured at a given 
site represents a summation of the appar- 
ent resistivities to a depth approxi- 
mately equal to the electrode spacing 
and located between the two potential 
electrodes. 

The Wenner method is used in two ways. 
In the first the electrode spacing is 
held constant and the entire array is 
traversed across the ground surface. 
This results in looking at a large area 
to a constant depth. This constant-depth 
traverse may be used to look for geologic 
faults, shallow voids, lithologic con- 
tacts, etc. 

The second method is to maintain a con- 
stant center to the electrode array and 
incrementally increase the electrode 
spacings. This results in a deeper and 
deeper measurement depth with an asso- 
ciated larger and larger measurement 
volume. From this, data sounding curves 
may be constructed for interpretation of 
the earth materials at depth by use of a 

C, P| P2 C2 

I I I I Ground surface 



C-C, Cur 



FIGURE A-1. - Wenner array electrode configuration. 




FIGURE A-2. - Approximate current and poten- 
tial fields for Wenner array. 



method such as shown in figure A-3 or by 
means of the sounding curves developed by 
Mooney and Orelleno (6^). The volume of 
earth material expands as the cube of the 
electrode spacing, which tends to have an 
averaging effect on irregular subsurfaces 
and to obscure small features at greater 
depths. 

The averaging effect is useful when 
trying to determine the properties of 
large volumes of material, such as mea- 
suring the fracture density in an ore 
body or measuring solution fronts for in 
situ mining. 

The equation for calculating Wenner 
apparent resistivity is 



AV 



'app 



k apparent resistivity- 



ohm-meters, 



where AV = potential difference in volts 
between P, and P2. 

I = current injected between C^ 
and C2 to create potential, 
in amperes. 

k = a geometric constant. 

a = electrode spacing, meters. 

k = 2TTa (0.3048 if a is in feet). 

(k = 3ira to a 4Tra for underground 
measurements. ) 



1,000 



"a" 






spacing, 


Pa 


2^Q 


m 






0.3 


1 1 5 


1 15 


.6 


65 


180 


.9 


42 


222 


1.2 


64 


286 


1.5 


56 


342 


1.8 


168 


510 


2.1 


132 


642 


2.4 


168 


810 


2.7 


145 


995 . 


3.0 


160 


1,115 vS 




• ^ Inferred layer depth at 
I inflection point of curve 



"a" SPACING, m 

FIGURE A-3. '■ Moore cumulative method of 
depth interpretation. 



APPENDIX B.— DESCRIPTION AND INTERPRETATION OF POLE-DIPOLE METHOD 



35 



The pole-dipole electrode configuration 
used by the Bureau Is a modification of 
the method developed by the English spe- 
leologist, Brlstow, for detecting caves. 
It uses several current electrodes, one 
at a time, near the measurement site and 
the other at effective Infinity (5 to 10 
times the largest separation between the 
other current electrode and the potential 
electrodes). The potential electrodes 
are moved as a pair with a constant, pre- 
determined separation between them. 
Figure B-1 Illustrates the electrode con- 
figuration used and the electrical fields 
developed by this method. As may be seen 
by comparing this figure with figure A-2, 
the volume of material being measured for 
a given depth of Investigation Is much 
smaller than for the Wenner; this results 
In a much higher detection resolution for 
a given sized target at a given depth. 



possible as with the Wenner method. 
Figure B-2 shows a few of the measurement 
fields developed during this procedure 
using the automated resistivity system. 
As may be seen when comparing the poten- 
tial fields from numerous current elec- 
trode positions, a high degree of data 
redundancy Is created. Increasing the 
accuracy of the method. 

The equation for calculating pole- 
dlpole apparent resistivity Is, in 
ohm-meters , 



where 



2tt 


AV 

I 




Papp I I 
ri T2 




= distance C^ - 


Pl. 


in meters. 


= distance C^ - 


P2. 


in meters. 



The method may be used for performing 
either a constant-depth traverse or a 
depth sounding, but in practice a com- 
bination of the two is run, resulting in 
a cross-sectional view of the earth below 
the array position. The amount of data 
resulting from this cross-sectioning 
is so vast that the interpretation is 
done by computer. For this reason, a 
field interpretation of the data is not 



AV = voltage potential measured 
between P^ and P2. 

I = current injected at C^ to 
create potential, in 
amperes . 




FIGURE B-1. - Approximate current and poten- 
tial fields for pole-dipole (Bris- 
tow) method. 



FIGURE B-2. 



A few of the measurement fields devel 
oped by the pole-dipole method. 



36 



ELECTROMAGNETIC GROUND RADAR METHODS 
By Richard J. Leckenby'' 



ABSTRACT 



The Bureau of Mines has developed a 
number of electromagnetic ground probing 
radar techniques for use by the mining 
industry to detect and map potential min- 
ing hazards , such as abandoned mines , 



water-filled fractures and faults, veins, 
and well casings. This paper gives in- 
sight to some of the techniques and how 
they are used for detection and mapping 
of various hazards. 



INTRODUCTION 



The Bureau of Mines has been actively 
involved in electromagnetic ground prob- 
ing radar methods for over 6 years. In 
that time frame, the research effort has 
evolved from experiments proving the fun- 
damentals and capabilities of the tech- 
nique to the development of better detec- 
tion systems with improved processing and 
display capabilities that will be useful 
and practical to the mining industries. 
The Bureau has actively pursued research 
and development in a number of electro- 
magnetic methods, each having its own 
unique advantages and disadvantages. 



In general, it is believed that elec- 
tromagnetic techniques can or will pro- 
vide the mining industry a tool for rapid 
and undisruptive detection and mapping of 
geological features and potential haz- 
ards, such as abandoned mines, faults, 
fractures, veins, and well casings. The 
intent of this paper is to present an 
overview of some of the electromagnetic 
techniques the Bureau has and is develop- 
ing, and how these techniques are used 
for the detection and mapping of poten- 
tial hazards. 



ACKNOWLEDGMENTS 



Richard L. Myers and James J. Snod- 
grass, geophysicists at the Bureau's Den- 
ver Research Center, are acknowledged 
for their work as Principal Investigators 



and Technical Project Officers, for the 
development of many of the electromag- 
netic techniques discussed in this paper. 



DISCUSSION OF ELECTROMAGNETIC GROUND PROBING RADAR METHODS 



The term electromagnetic is a broad, 
encompassing term. The first priority in 
discussing any electromagnetic technique 
is to define it by placing limits on the 
technique to be used. In the case of the 
techniques to be discussed here, the 
methods will be confined to the use of 
electromagnetic radiation for the mea- 
surement of the electrical and magnetic 
properties of the ground as a hazard 
detection method. Further restrictions 
will be placed to include only that band 
of frequencies in the electromagnetic 



Physicist, Denver Research Center, Bu- 



reau of Mines, Denver, Colo. 



spectrum that can provide adequate pene- 
tration and resolution. In order to 
provide this adequate penetration and 
resolution in a variety of minable mate- 
rials, frequencies between 5 and 500 MHz 
are normally used. This translates to 
wavelengths between 60 and 0.6 m in free 
space, and 20 to 0.2 m in a ground media 
having a velocity of propagation approxi- 
mately one-third that of free space. 
Using this band of frequencies, penetra- 
tion from 2 to 300 m can be achieved for 
most rock types, for a large number of 
rock media where the technique would be 
useful, penetration of 30 m or better has 
been or soon will be achievable. 



37 



The underlying principle for electro- 
magnetic techniques is to determine and 
measure the effects the ground has on the 
electromagnetic radiation as it passes 
through the ground. The major physical 
parameters affecting the propagation are 
the frequencies of the waves, the complex 
permittivity, and the complex permeabil- 
ity of the ground. Whenever there is a 
change in either the complex permittivity 
or permeability, the electromagnetic wave 
propagation characteristics will change. 
That is, reflection or refraction will 
occur with associated changes in veloc- 
ity, amplitude, and polarization taking 
place for a wave of a given frequency. 
It is these changes that are measured, 
and with general geological knowledge of 
an area, an interpretation can normally 
be made, assigning a given wave propaga- 
tion change to a probable geological 
feature. 

Although there is a variety of differ- 
ent methods, techniques, and instrumenta- 
tion associated with the various electro- 
magnetic methods, there are also some 
commonalities. For instrumentation, fig- 
ure 1 illustrates the key components 
found in most of the electromagnetic 
ground probing radar techniques. Like- 
wise, most of the methods are used in 
either a reflection (radar) type or a 
transillumination (one-way) type mode. 
Figure 2 is illustrative of the two 
modes, where in this case the anomaly, 
represented as an abandoned mine, has a 
different electrical property than the 



Transmitting 
antenna 



Receiving 
antenna 





Transmitter 



Receiver 



Controller 



Display 



FIGURE 1. - Key components for most ground radar 
systems. The linking of the transmitter 
to the controller is optional depending 
on the technique used. 

surrounding rock material. This change 
causes reflection and refraction to 
occur. The reflection, the echo, or the 
refracted wave is received via an antenna 
and then is recorded. It should be noted 
that for the radar shown where two an- 
tennas are used, a surface and an air 
wave are received, as well as an echo. 
For single-antenna systems, the surface 
wave would not be present. 



38 



Transillumination 
(One-way) 




FIGURE 2. - Two modes of operation: reflection and transillumination. The solid black lines with the 
arrows indicate named travel paths of electromagnetic waves from the transmitter to re- 
ceivers, e^ and o] are the dielectric constant and conductivity of the rock media. Thedi- 
electric constant and conductivity of the anomaly are represented by £3 and a2« The link 
between the transmitter and receiver is optional depending on the technique used. 



Some of the electromagnetic methods 
that can be used are (1) short-pulse, 
(2) synthetic-pulse, (3) continuous-wave 
(CW), (4) FM-CW, (5) chirp, and (6) tone 
burst. They all have their own advan- 
tages and disadvantages, unique electron- 
ics, data processing, and display re- 
quirements. From this list, the Bureau 
has mainly been concentrating its efforts 
toward developing and researching 
the short-pulse, synthetic-pulse, and 
continuous-wave methods. 



the ground surface, a borehole, or under- 
ground in a mine. The method of detec- 
tion is usually done by making voltage 
versus time measurements, using a sam- 
pling oscilloscope or high-speed transi- 
ent digitizer. The measured voltage is 
then recorded either on digital or analog 
tape. In the radar mode, the anomaly is 
detected by receiving the echo of the 
transmitted wide-frequency band pulse at 
a delayed time. 



Figure 3 illustrates the short-pulse 
method. The method can be applied from 



39 



Transmitting 
antenna 



(D 



Receiving 
antenna 



(b 



II ■ w V 



Transmitted 



Anomaly 




One-way 



G 



Receiving 
antenna 



(Amplitudes scaled for presentation) 



FIGURE 3. - Short pulse radar methods. Representative pulse signals with respect to time, 
t, of the transmitted and received signals are shown for different travel paths 
indicated by the arrows for both radar and one-way modes. 



For the one-way mode, the signal will 
arrive at a different time and amplitude 
than what would be expected If there were 
no anomaly; thus, an anomaly within the 
ground can be detected. Some of the 
advantages of the short-pulse method are 
that (1) the Instrumentation can be 
assembled using standard off-the-shelf 
equipment, (2) preliminary results can be 
quickly displayed, (3) data processing 
techniques have already been developed 
for a variety of signal problems, and (4) 
Interpretation of the data Is possible 
without elaborate processing or display 
techniques. 

The continuous-wave method Is shown In 
figure 4. The name Is Indicative of the 
technique, in that a continuous wave 
is transmitted into the ground. The 



technique is normally used in a transil- 
lumination mode, and the Bureau uses the 
method mainly from cross-borehole sur- 
veys. A measurement of the amplitude and 
phase of the received signal is made and 
usually referenced to the transmitted 
signal. Some of the advantages of the CW 
method are that the electronics can be 
designed for a given frequency being 
transmitted, and thus can be made more 
efficient than those used in the short- 
pulse method. The recorded data can also 
lend themselves to tomographic-type dis- 
plays, making them easier for interpreta- 
tion. The disadvantage of the technique 
is that it normally has to be applied in 
a transillumination mode, because the 
return signal in the radar mode is so 
confused that it is Impractical to inter- 
pret the data. 



40 



Radar 



Transmitter 



Controller 




Receiver 



One -way 



t > 



vvv\. 



A. COS(wnt) 



Anoma ly 




Aq cos(wot + 4>q) + Ag cos(cuot + 4>^) + Ar cos (uiqI + <l>r) 



A, cos(wot + </),) 
I 1 > 



\jj~ Receiv 



FIGURE 4t - Continuous wove methods. Representative continuous wave signals with 
respect to time, t, for the radar and one-way mode. A^, A^, A^, and A ^ are 
ampi itudes of the continuous waves with on angular frequency of a^ for the 
transmitted surface, reflected, and refracted waves, respectively. cf)^,ct),, 
and are phase constants of the surface, reflected, and refracted waves. 



The synthetic-pulse method is an 
attempt to take advantage of the posi- 
tive points from the CW and short-pulse 
methods. The electronics for trans- 
mitting and receiving the signals are 
somewhat similar to that used for con- 
but for 



tmuous waves, 
method, hundreds of 
are transmitted and 
cessing the return 



the synthetic 
discrete frequencies 
recorded. By pro- 
signals, pulselike 



data can be constructed. The synthetic- 
pulse method can be used in a radar mode, 
and at this time it shows promise of at 
least doubling the effective range of the 
short-pulse method. 



Although all of the different methods 
can be used for detection of the various 
anomalies, locating or mapping the poten- 
tial hazards or geological features is 
necessary in order to make the methods 
practical. The mapping is performed nor- 
mally by moving the antenna or antennas 
relative to the anomaly, and thus causing 
a change in travel distance for the sig- 
nals . This movement can be performed by 
a number of ways . Some of the more com- 
mon ways are known as (1) constant off- 
set; (2) moveout or common depth; and (3) 
cross-borehole mapping. 



41 



The constant offset (fig. 5) is com- 
monly used for the radar mode. The 
transmitting and receiving antennas are 
moved together with a constant separation 
distance. This technique can be applied 



on the ground surface, in a borehole, or 
along a wall or a working face in an 
underground mine. The resulting data 
appear similar to that in figure 6. 



probe .V . 




FIGURE 5. - Constant offset mapping techniques. The transmitter and receiver are separated 
by a constant distance, X, and are moved together along the surface or borehole 
to provide for change in travel distances between the transmitter-receiver and 
the anomaly. 



42 



100 



300 - 



400 




-17.5 -15.5 -13.5 -11.5 

STATION LOCATION, m 



FIGURE 6. - Exampleof data using aconstant offset technique. For each transmitter-receiver 
station location, a short pulse wave form is recorded v/ith respect to time. The 
first pulse shape after time zero is the air-surface wave. Pulses recorded later 
are reflection related. A time of 400 nsec, in this case, represents a depth of 
around 30 m. 



43 



The conunon depth point (CDP) mapping 
technique is shown in figure 7. This 
technique maps an anomaly by separating 
the transmitting antenna from the receiv- 
ing antenna. The target or anomaly 
remains fixed, but the travel distance 
between the antennas increases with each 
antenna movement. Figure 8 illustrates 



the CDP approach as it was applied in a 
coal mine using the synthetic-pulse 
method. Using the change in arrival 
times for the reflections, and knowing or 
estimating the velocity, it is possible 
to determine the distance of the anomaly 
from the transmitting network. 



Controller 



Surfoce^^ 



Receiver / \Transmitter 




FIGURE 7. - Common depth point mapping technique. The transmitter and receiver are 
moved apart from each other, causing a change in travel paths for each 
reading. 



44 




FIGURE 8. - Reflection in coal using common depth 
mappingt Pulses occurring around 3 
^tsec are reflection through 50 ft of coal. 
The slope of the reflections can be used 
to determine the di stance of the reflector 
if the velocity is known. 

For cross-borehole mapping, the trans- 
mitting and receiving antennas are placed 
in different boreholes. By moving the 
two relative to each other, a cross-hatch 
pattern is possible, as shown in fig- 
ure 9. From this pattern, a recon- 
structed image of the electrical proper- 
ties of the ground can be made. 

SUMMARY 

The intent of this paper is to intro- 
duce electromagnetic ground probing radar 



methods to those unfamiliar with the 
technique. This was done by showing some 
of the different methods and how they are 
applied in detecting and mapping geologi- 
cal features and potential hazards. The 
Bureau has now conpleted feasibility 
tests of short-pulse surface and borehole 
techniques, synthetic-pulse methods for 
detecting hazards in coal mines, and 
cross-borehole continuous-wave studies. 
In all cases, the results have been en- 
couraging, and prototypes are now in the 
design or testing stages for each method. 
Within the next 2 years, the results of 
the prototype systems should be com- 
pleted. With these future tests, it is 
hoped that performance figures can be 
established, so that a better understand- 
ing of what anomalies can be determined, 
at what distance, and in what media can 
be known to the potential users. As of 
now, it is still tricky to predict the 
results of the various methods and tech- 
niques, and the use of any of the methods 
should still be approached as being pos- 
sible and promising, but not conqjletely 
proven. 



Borehole No. 



v:jv;:::;;"-JyF^ 



Transmitters 




FIGURE 9. - Cross borehole mapping technique. By moving the transmitter and receiver to vari- 
ous locations, a variety of travel paths are possible- By comparing the different 
received signals, detailed analysis of the material betvv'een holes is possible. 



BIBLIOGRAPHY 2 



45 



1. Belsher, D. R. Detection of Lost 
Oil Well Casings and Unknown Water-Filled 
Voids in Coal Mines Through Development 
of a Microwave Antenna System. BuMines 
Open File Rept. 6-79, February 1978, 
94 pp.; contract H0272007, National Bu- 
reau of Standards. 

2. Cook, J. C. Radar Transparencies 
of Mine and Tunnel Rocks. Geophysics, 
V. 40, No. 5, October 1975, pp. 865-885. 

3. Dines, K. A., and R. J. Lytle. 
Interactive Reconstruction of Underground 
Refractive Index Distribution From Cross- 
Borehole Transmission Data. Lawrence 
Livermore Lab. , Tech. Rept. UCRL-52348, 
November 1977, 6 pp. 



4. Fowler, 
. T. Houck. 



J. C, S. D. Hale, and 
Coal Mine Hazard Detection 



•^ Items ^, 4, and 7 are available for 
reference at the Denver Research Center, 
Bureau of Mines, Denver, Colo. 



Using Synthetic Pulse Radar. BuMines 
Open File Rept. 79-81, January 1981, 
84 pp.; contract H0292025, ENSCO, Inc. 

5. Kracchman, M. B, Handbook of Elec- 
tromagnetic Propagation in Conducting 
Media. U.S. Navy, Naval Material Com- 
mand, NAVMAT P-2302, 1970, 128 pp. 

6. Okada, J. J., E. F. Laine, R. J. 
Lytle, and W. D. Daily. Geotomography 
Applied at the Stripa Mine in Sweden. 
Lawrence Livermore Lab. , Tech. Rept. 
UCRL-52961, April 1980, 24 pp. 

7. Suhler, S. A., and T. E. Owen. 
Development of Deep-Penetrating Borehole 
Geophysical Technique for Predicting Haz- 
ards Ahead of Coal Mining. BioMines Open 
File Rept. 77-80, October 1976, 110 pp.; 
contract H0252033, Southwest Res. Inst.; 
available from National Technical In- 
formation Service, Springfield, Va. , 
PB 80-208614. 



46 



IN SITU NEUTRON ACTIVATION ANALYSIS 
By George J. Schneider! 



1 



ABSTRACT 



An in situ neutron activation analysis 
system has been developed and tested in 
several mineral deposits. The 2-in-diam 
borehole logging sonde in the system con- 
tains a microprocessor-controlled, high- 
resolution, gamma-ray spectrometer with 



digital data transmission to the surface 
on 4-H-O logging cable. Californium is 
used as a neutron source for activation. 
The gamma-ray spectrometer has been 
adapted to commercial service for assay- 
ing uranium ores in disequilibrium. 



INTRODUCTION 



In September 1976, after evaluating 
con5)etitive proposals, the Bureau of 
Mines let a contract^ to Princeton 
Gamma-Tech, Inc. , to design a unique 
borehole assaying system. Their research 
has resulted in the development and com- 
mercial availability of a logging system 
that still defines the state-of-the-art 
for in-place ore grade analysis. The 
logging sonde includes a high-resolution 
intrinsic germanium detector cooled by a 
solid or melting cryogen; a 4,000- 
channel, microprocessor-controlled, mul- 
tichannel analyzer; and a digital com- 
munications link, along industry standard 
4-H-O logging cable to a microcon^juter- 
based data storage and display system at 
the surface. Small quantities of 
the man-made isotope, calif ornium-252, 
aew used as a source of neutrons for 
activation. 

The advantages of downhole data pro- 
cessing are demonstrated by the measured 



2 keV resolution at 1.33 MeV, of the 
borehole gamma-ray spectrometer over 
3,000 ft of 4-H-O logging cable. Previ- 
ous systems have only provided about 
14 keV resolution over 1,000 ft of coaxi- 
al cable at the same energy. The practi- 
cal benefit is virtually laboratory- 
quality gamma-ray spectra for in situ 
analysis. Figure 1 is a coti5)lete pronpt 
capture gamma-ray spectrum from a 
magnetite deposit acquired by the sonde 
with a 2.2-yg calif ornium-252 source in 
a 5-in-diam borehole; it illustrates 
the high-resolution spectra routinely 
obtained. 

The reliability of the system is demon- 
strated by the daily commercial use of 
derivative systems for assaying uranixim 
ores. The contractor has operated sever- 
al systems in regular service for nearly 
2 years. A derivative of the multichan- 
nel analyzer design has also been adapted 
to a commercial laboratory analyzer. 



NEUTRON ACTIVATION ANALYSIS 



Neutron activation analysis is an in- 
herently effective technique for borehole 
assaying. The method is, in general, ab- 
solute and volumetric, and it exceeds 
mining industry requirements for elemen- 
tal sensitivity in near-real-time analy- 
sis. A source of neutrons, californium- 
252 in the present Bureau system, 

! Geologist, Denver Research Center, Bu- 
reau of Mines, Denver, Colo. 

^Contract H0262045, "A Borehole Probe 
for In Situ Neutron Activation Analysis." 



irradiates the rock surrounding an access 
borehole. The neutrons are slowed to 
thermal equilibrium with the rock mass 
from an initial energy between 1 MeV and 
8 MeV by many elastic and inelastic col- 
lisions with the nuclei of individual 
atoms in the rock. Elastic scattering 
describes a collision of a fast neutron 
and an atomic nucleus, where kinetic 
energy is conserved. In inelastic scat- 
tering, a net loss of kinetic energy 
occurs. An inelastic collision raises 
the energy of the atomic nucleus to an 



47 




" 400 - 




f2 300 



F(7632;7646') 



Fe(7632,764 6) 



^\J 



^^'^^^^^^'^^ 



9.5 



FIGURE 1. = In situ gamma ray spectra from magnetite (iron ore) deposit near Dover, NJ- This 
is a complete 4,000-channel prompt-capture gamma ray spectrum obtained with the 
logging system (in four parts). 



unstable state and the nucleus returns to 
a stable state through the emission of 
ganima rays at discrete energies. 

After losing energy in many collisions, 
a neutron reaches thermal equilibrium 
with the rock mass. Thermal neutrons are 
captured by atomic nuclei in the rock 
according to well defined probability or 
cross section for each isotope. Neutron 
capture raises the energy of a nucleus to 
an unstable state. With only a few ex- 
ceptions all the naturally occurring iso- 
topes return to a stable state by the 
emission of one or more gamma rays at 
discrete energies; and in decay reac- 
tions, by the emission of another sub- 
atomic particle. An individual nucleus 
can return to a stable state by only one, 



or one sequence of, several available 
reaction paths, but a group of nuclei of 
the same isotope will follow all avail- 
able reaction paths according to very 
precisely known ratios. 

Prompt-capture gamma rays are produced 
by the immediate release of energy after 
thermal neutron capture. Decay gamma 
rays are produced by reactions with a 
specific statistical half life. Thermal 
neutron activation, both prompt capture 
and decay, is in practical terms the most 
important class of neutron activation re- 
actions for in situ analysis, because 
these reactions have the highest cross 
section or probability of occurrence for 
most isotopes and are the most prolific 
producers of gamma rays. 



48 



In the earth, the rate of kinetic 
energy loss by a neutron, from a median 
energy of about 2.5 MeV when emitted by 
spontaneous fission from a californium- 
252 source to about 0.025 eV at thermal 
equilibrium, depends to a first approxi- 
mation, on the amount of water or por- 
osity in the rock. The hydrogen nucleus 
has about the same mass as a neutron, and 
energy transfer from the neutron to the 
nucleus is efficient. In porous satur- 
ated rock, neutrons are thermalized in a 
smaller volume than in dry media, and 
consequently, the efficiency of detecting 
gamma-ray spectra from thermal neutron 
activation increases. In tests of the 
Bureau system, a ratio in proportion to 
the square root of the number of counts 



in the 2.23 MeV prompt-capture gamma ray 
from hydrogen was found to be an effec- 
tive correction for changes in porosity 
of rock. 

The volume of investigation for in situ 
neutron activation analysis is determined 
by the energies of the gamma rays logged, 
and in general the radius of investiga- 
tion increases with gamma-ray energy. A 
gamma ray with an energy of 500 keV will 
penetrate about 5 cm of rock of average 
density, while a gamma ray of 5 MeV will 
penetrate about 40 cm of rock. In gen- 
eral terms, most elements produce prompt- 
capture gamma rays above 3 MeV in energy, 
and useful decay gamma ray peaks between 
400 keV and 3 MeV. 



THE LOGGING SYSTEM 



The borehole sonde (fig. 2) designed by 
Princeton Gamma-Tech, Inc., for the Bu- 
reau has a 2-in diameter. The detector 
and cryostat are in one section 4 ft 



--4-H-0 logging cable 
Cable head 




"7 "Power supplies 



jT-'Asynchronous serial 
line interface 

-4,000-channel, microprocessor- 
based multichannel analyzer 
- Amplifier 



Cryostat 



-- P type germanium 

gamma-ray detector 



^~- Shadow shield 
r-- Californium source 



IXl ^ '2 inches diameter 
FIGURE 2. - The logging sonde. 



long, and the electronics are in a sepa- 
rate section 6 ft long. A tungsten shad- 
ow shield and nylon spacers were config- 
ured in various lengths between 18 in and 
5 ft during specific tests to separate 
the californium source from the gamma-ray 
detector. 

Several Freon and propane compounds 
frozen by liquid nitrogen and, during one 
test, a solid slug of copper cooled to 
77 K have been used to cool the intrinsic 
germanium detector. The cryogens are 
contained in removable canisters that can 
be inserted and removed from the sonde in 
a few minutes. A single canister of 
Freon-22 permits over 8 hr of logging. 

The electronics package in the sonde 
contains a preamplifier, an amplifier, a 
high-voltage supply for detector bias, a 
successive approximation type analog- 
to-digital (ADC) converter, low-voltage 
power supplies for the components, a 
Motorola 6800 microprocessor with 4,096 
eight-bit words of memory, and a half du- 
plex serial cable transceiver. A total 
of 3,968 words of memory were used for 
data storage. Conversion time, indepen- 
dent of pulse height, was 16.6 msec for 
the ADC. 

The operation of the system is shown 
schematically in figure 3. The intrinsic 



49 



Data 

terminal 

("T.I." 743 KSR) 



P.G.T. 385 E 
Dual floppy 



(Computer interface) 



Fast serial input 
Slow serial output 



Incremental en- 
coder up-down 

Distance counte 



Drum controller 



Bidirectional 
incremental 
stioft encoder 




Motor 
Power driver 



(MOBILE INSTALLATION) ] Line receivers-driverr] 
SURFACE 



Cable drum 
motor speed 



Unregulated 
power supply 



DOWN HOLE 
" SONDE " I Line drive-receive 



I Power supply regulators 



Fost serial output 
Slow serial input 



Program 
ROM 



MircD- 
rocessor I 



Data 

nicroorocessor 
RAM 



High. voltage 
bios supply 



Peript)eral 

Interface 

A 



Periptieral 

Interface 

B 



Peak 
detector-hold 



Averaging 
S-A A.D.C- 



Preamp-fixed 
gain amp 



Ge 
detecto 



Source 



Gain stgb. 
register d/A 



Zero stab, 
register D/A 



FIGURE 3. ■= Operating logic for the logging system. 



germanium detector converts the energy of 
a gamma ray within the detector to a 
voltage pulse proportional in amplitude 
to the energy of the incident gamma ray. 
The voltage signal is filtered, shaped, 
and amplified before input to the ADC. A 
digital number equivalent to the ampli- 
tude of the voltage pulse is transmitted 
from the ADC to the microprocessor. The 
microprocessor stores the number from the 
ADC as a single count in an appropriate 
location in memory corresponding to the 
energy of the incident gamma ray. Data 
are accumulated and stored in memory for 
several seconds at an effective rate of 
10,000 to 20,000 counts per second, and 
then transmitted along 4-H-O logging 
cable to a microcomputer at the surface. 
The microprocessor transmits the number 
of counts in each channel in succession 
and resets each channel back to zero. 
Transmission is interrupted by any new 
event from the ADC and resumes after the 
incoming event is processed and stored in 



memory. Data transmission is over a 
single wire of four-conductor 4-H-O ca- 
ble at a rate of 31,800 bits per second 
(31.8 baud). 

A microcomputer at the surface receives 
the data at a serial line interface off 
the logging cable. The data are stored 
on magnetic tape or disks, and processed 
for display on a terminal in the logging 
van whenever convenient for the operator. 
The microcomputer at the surface controls 
the cable which through a digital- 
to-analog converter (DAC). The circuitry 
and components for the microprocessor 
based multichannel analyzer, preamplifi- 
er, amplifier, and power supplies in the 
sonde are available in the previously re- 
leased Phase I project report at Bureau 
libraries . 

Revised circuit diagrams for the ADC 
are shown in figure 4, and for the micro- 
processor-based analyzer in figure 5. 



50 



POWER RESE 




NOTE 

* =aN/>LOG GND 
A =POWEB GND 



FIGURE 4, - Circuitry and components for the downhole ADC. 




Iter; ^jaiijifiHfiShfrHr.. II ■.;:>- -r^ 



FIGURE 5. - Circuitry and components for the microprocessor-based downhole multichannel analyzer. 



51 



FIELD TESTS 



Uranium 

The first field test of the system was 
in disequilibrium uranium ores near San 
Antonio, Tex., in cooperation with the 
U.S. Department of Energy and Continental 
Oil Co. (CONOCO). Because uranium is 
naturally radioactive, no neutron source 
was necessary, and the high-resolution 
gamma -ray spectrometer could be evaluated 
without complication. 

Uranium ores contain uranium-238 and 
uranium-235 with their decay products or 
daughters , thorium with its daughters , 
and potassium-40. The proportion of 
uranium-235 to uranium-238 is accepted as 
a fixed ratio. Gross gamma and KUT (po- 
tassium, uranium, thorium) logs estimate 
the amount of uranium-238 present by mea- 
suring the activity of one of the daugh- 
ters, bismuth-214. In ideal geologic en- 
vironments, bismuth-214 is present in 
equilibrium to uranium-238, and the gross 
gamma or KUT log is valid. But in many 
real enyironments disequilibrium is com- 
mon between uranium and the long- 
half-life daughters because of differ- 
ential leaching or other mechanisms for 
removal of uranium or its daughters se- 
lectively. The measurement of equivalent 
uranium by conventional methods in these 
environments is in error since the con- 
centration of uranium and bismuth-214 are 
not in proportion. In south Texas and 
parts of New Mexico and Wyoming, dis- 
equilibrium ores are typical, requiring 
expensive core drilling for economic 
evaluation of the deposits and mine 
planning. 

Two daughters of uranium-238 establish 
equilibrium with the parent isotope in a 
few months and are therefore always in 
equilibrium under geologic conditions. 
The first daughter of uranium-238, 
thorium-234, has a half life of 24.1 
days and decays to protactinium-234. 
Protactinium-234 decays with a 1.17-min 
half -life, accompanied by the emission of 
a gamma ray at 1.001 MeV. Because the 
gamma-ray is produced only 0.59 pet of 
the time an atom of uranium-238 decays. 



and is also close to a major spectral 
peak of bismuth-214, the protactinium 
gamma ray cannot be logged with conven- 
tional systems. The high resolution in- 
herent in the Bureau system allows the 
protactinium peak to be discriminated 
from adjacent peaks. 

Thirty boreholes representative of sev- 
eral uranium-bearing lithologies under 
both positive and negative disequilibrium 
and equilibrium were logged during the 
field test. Figure 6 shows the results 



600 



500 



400 



300 



200 



100 



1 r 



^ Closed can 

!E!3 Gross gamma log 

'' ■ Pa-234 log 

• Labchem fluorometric 
and colorimetric 




100 



FIGURE 6t - Laboratory and in situ analysis in 
Texas uranium ore. The Pa-234 (protactin- 
ium) log provides a correct in situ assay 
through an ore zone where uranium is in dis- 
equilibrium with the long-half-life daughters. 



52 



of gross gamma logging confirmed by 
closed can laboratory analysis of core 
sample, and the protactinium (Pa-234) log 
confirmed by beta minus gamma (B-y), col- 
orimetric, and f luoriometric laboratory 
assays of core samples from a test bore- 
hole in Texas uranium ore. Note the top 
of the ore zone is in equilibrium. Ura- 
nium has been leached from the center of 
the ore zone and redeposited immediately 
below the leached zone. The Pa-234 log 
shows the distribution of uranium cor- 
rectly, while the gross gamma log does 
not. The tests conclusively demonstrated 
the feasibility of the technique. The 
contractor now offers the method as a 
commercial logging service with wide 
acceptance. 



gamma ray spectrum was calibrated over 
the region including the 657.7 keV 
silver-llOg gamma ray by fixing the loca- 
tion of the 511-k.eV positron annihilation 
peak and the 1,778-keV decay peak from 
aluminura-28. The 2,223-keV pronpt cap- 
ture gamma ray from hydrogen was used to 
normalize the data for formation poros- 
ity. The foot-by-foot data were averaged 
into assays representative of 5 ft-long 
intervals to compare with laboratory 
assays on samples from the drill holes. 
The good agreement between the in situ 
analysis and the laboratory assays is in- 
dicated in figure 7, the results from 
logging one of the test boreholes. 

Other Field Tests 



Silver 

Naturally occurring silver metal con- 
sists of two isotopes, silver-107 and 
silver-109. Silver-109, 48.18 pet of 
natural silver, activates to produce two 
isomers of silver-110 by thermal neutron 
capture. The reaction producing silver- 
llOg, the ground state, has a large prob- 
ability of occurrence or cross-section of 
88 barns. The reaction producing the 
second isomer, silver-llOm, a metastable 
state has a much smaller cross section of 
4.4 barns. Silver-llOg decays by beta 
emission and electron capture with a 
prominent (4.5 pet intensity) gamma ray 
at 657.7 keV with a half -life of 
24.4 sec. 



Field tests were also conducted in iron 
ores in cooperation with Halecrest Mining 
Co., copper porphyry ore in cooperation 
with Kennecott Corp. , metallurgical coal 
in cooperation with the United States 
Steel Company and the U.S. Geological 
Survey, and gold ore in cooperation with 
Golden Cycle Mining Co., Texasgulf Corp., 
and the U.S. Geological Survey. The re- 
sults of these tests will be discussed in 
detail in future publications. 



CRC-14 
Chemical assay of cuttings 



The large cross-section and short half- 
life indicate a reaction particularly 
suitable for in situ analysis. 

In cooperation with Chevron Oil Field 
Research, Chevron Resources (CRC), and 
Minerals Engineering Co. , the system was 
field tested in silver bearing ore near 
Creede, Colo. The procedure adopted for 
the test was to irradiate a 1-ft region 
in the borehole for 1 min, about 
2.5 half-life for the silver-llOg re- 
action, and then to move the germanium 
detector into the region for 1-min to 
measure the activity of silver-llOg. The 



Scale tor solid line-raw 
Scale for dashed 



average 

ne-H normalized 



Delayed / N-activation on silver 




L 



^ 



25 

20 o 

15 - 
'a 

5 < 



70 90 110 130 150 170 190 21 
DEPTH, ft 



FIGURE 7. - Compilation of 5-ft average silver assay 

values from foot-by-foot measurements. I 

The effect of the hydrogen correction for 
porosity is also shown. 



FUTURE RESEARCH 



53 



Further development of borehole neutron 
activation analysis is planned by replac- 
ing the californium sources in the system 
described with a pulsed, sealed-tube neu- 
tron generator, and a small linear accel- 
erator. The use of an accelerator as a 
neutron source eliminated the shielding 
required to protect operating personnel, 
since neutrons are produced only when 
potential, typically 100 keV, is applied 
to the target in the generator. Sensi- 
tivity will improve for decay gamma-ray 
logging because of increased neutron 



flux, and for prompt-capture gamma-ray 
logging because the spectra are collected 
between pulses in much improved back- 
ground. Also, inelastic scattering re- 
actions will be usable because of the 
neutron energy available with an acceler- 
ator type source. Figures 8 and 9 show 
the estimated sensitivity obtainable for 
most elements of economic interest with a 
fully developed system by in situ neutron 
activation analysis and other related in 
situ assaying methods within real opera- 
tional constraints. 



*h' 




BOREHOLE ASSAY PROGRAM 










He 


Li, 


Be^ 




*B* 


< 


1 
N 


^0- 


F^ 

• 


Ne 


f A 

Na 


Mg^ 


Al- 

• 


• 


P^ 


• 


*C1* 


f 
Ar 


f A 

■K 




f 
Sc 




^ 


• 


o • 


o • • 


Vo 


• 


•Cu* 

o •• 


Zn 

o •• 


Go 


Ge 


a's 


sV 


B'r 


k\ 


Rb 


▼ 

Sr 

• 


Y 

• 




Nb 

• 


Mo 

• 


Tc 


Ru 

• 


▼ 

Rh 


Pd 


Ag 


▼ A 

Cd 

• 


In 


«n 

• 


f 

Sb 

o • 


Te 

• 


1 


Xe 


Cs 


Bo 

0»* 


Va 


H'f 


Ta 

• 


? 
W 

o« 


Ve 


0*s 


Ir 


Pt 

• 


Vu 

• 


♦ f A 

Hg 

o .^ 


Tl 

• 


1^ 


Bi 

• 


Po 


At 


■Rn 


Fr 


f 

"Ra 


■Ac 

































Ce 


Pr 


Nd 


Pm 


Sm 


Eu 


Gd 


Tb 


Dy 


Ho 


Er 


Tm 


Yb 


Lu 


■Th 

• 


■Pa 


o • 


Np 


Pu 


Am 


Cm 


Bk 


Of 


Es 


Fm 


Md 


No 


Lr 



METHODS 

■ Natural radioacfive 
decay 

• X-ray fluorescence 
o Gamma-gamma 

* Photon-activation 



FIGURE 8. 



NUCLEAR 

♦ Neutron-neutron 
▼ Thermal-neutron 

activation 
A Prompt-capture 

neutron activation 

•^ Fast neutron inelastic 
scattering 

► Fast neutron 
activation 

In situ assaying methods for elements of economic interest. Several methods have been demon- 
strated OS applicable. Selection of a method depends on whether the assay is to be done in 
boreholes, at the working face, or in haulage and on the sensitivity required. The neutron- 
based methods are in the left column. 



54 



VH"' 




BOREHOLE ASSAY PROGRAM 










He 


;'Liy' 


^ Be^' 




" B / 




; N '- 


:0': 


; F ^ 


>Ne^ 


Na :'Mg^^ 

iKi-ba; 

Rb |Srj 


rx:::::-: 

|AIJ 


J:s(: 


m 


.Vs;5 


Icii 


Ar 


Sc 


']t\-} 




Cr^ 


:Mn 


'^Fe: 


ICoj 


: Ni; 


IXXX-X- 

|Cu| 


iZni 


EGa; 


eGgI 


|As;: 


|Se| 


Br' 


iKrI 


IZrl 


|Nb= 


iMo^ 


Tc 


|Ru^ 


iiRhji 


Sx^v-:-:-:-:- 
:•:•:•:•:•:■:•:•:• 


11 


sCd^ 


::v.-.-.-.-.v 


|Sni 


%Sbi 


Ifel 


1 [ 


[Xe] 


ICsslBai 


La 


Hf 


ITal 


\ W : 


Re 


fosi 


103 


|Pt| 


lEI 


|Hg] 


ITii 


|Pb| 


|Bi| 


fPoi 


lAtJ 


Rn 


Fr |Ra| 


Ac 

































Ce 


Pr 


Nd 


Pm 


Sm 


Eu 


Gd 


Tb 


Dy 


Ho 


Er 


Tm 


Yb 


Lu 


• Th 


Pa 


III 


Np 


Pu 


Am 


Cm 


Bk 


Of 


Es 


Fm 


Md 


No 


Lr 



SENSITIVITY 
^*i < 1 pot 



HUB > 1 pet 

^^ > 100 ppm 

^^s'>g:i > 1 ppm 

FIGURE 9. - Estimated sensitivity that can be expected with neutron-based methods for various 
elements of economic interest. 



The operating conditions would be var- 
ied for the suite of elements of interest 
in a mineral deposit, but typical logging 
rates for the sensitivity indicated are a 
few feet per minute and always less than 
10 min per assay station with one 
exception. In situ gold assaying will 



probably require a two-step irradiation 
and spectra collection technique sepa- 
rated by about 8 hr for the sensitivity 
predicted. Logging rates in each step of 
the in situ gold assaying process are 
anticipated to be a few minutes per foot. 



55 



DEVELOPMENT OF AN IN-HOLE REPLACEABLE DIAMOND CORE BIT SYSTEM 
By W. C. Larson,'' W. W. Svendsen,2 R. E. Cozad,^ and J. R. Hof fmelster-* 



ABSTRACT 



A one-piece diamond core drill bit has 
been developed that can be removed for 
inspection or replacement without pulling 
the drill rods from the hole. The re- 
placeable bit system consists of two 
basic subsystems: The removable core 
bit, and the down-hole equipment neces- 
sary to replace the bit. The replaceable 
bit is designed for use in a standard "N" 



size wireline core drilling system. The 
system has undergone laboratory and full- 
scale field testing and is currently 
undergoing additional long-term field 
testing and engineering evaluation by 
the Longyear Co., Minneapolis, Minn. 
This report describes the replaceable 
bit system and summarizes its current 
status. 



INTRODUCTION 



Diamond core drilling is one method of 
drilling used in mineral exploration 
that, while costly and time consuming, 
provides the best "hands-on" information 
regarding the type, quality, and composi- 
tion of the rock being drilled. Histor- 
ically, improvements in diamond core 
drilling methods have developed very 
slowly, and most advances in core drill- 
ing have come, after years of need and 
long periods of development, as break- 
throughs in the state-of-the-art. The 
wireline core drilling system used in 
mineral exploration, now about 25 years 
old, is one example where significant 
time has been saved in drilling because 
the inner core barrel can be removed 
(raised and lowered on a wireline cable) 
from the drill string without pulling the 
drill rods out of the hole. 

Another major, time-consuming task in 
core drilling, estimated at between 5% 
and 10% of the total available working 
time (depending on hole depth), is the 
changing or inspection of the core bit. 

^Supervisory mining engineer. Twin Cit- 
ies Research Center, Bureau of Mines, 
Minneapolis, Minn. 

^Technical director, Longyear Co., 
Minneapolis, Minn. 

■^Project leader, Doerfer, Cedar Falls, 
Iowa. 

^Consultant, Minneapolis, Minn. 



As is well known in the drilling indus- 
try, withdrawing the drill rods from the 
hole is a laborious task. In addition, 
once the rods are removed, the drill hole 
is subject to caving owing to unstable 
rock formations. A logical advance, 
therefore, in the state-of-the-art of 
core drilling would be to incorporate the 
idea of replacing the drill bit without 
pulling the drill rods out of the hole 
(fig. 1). 

Many research organizations and drill- 
ing companies have recognized the advan- 
tages that could be realized by using a 
replaceable bit system. For example, 
domestic patents on a variety of retract- 
able bit concepts go back to the late 
1800' s and illustrate the long-time 
desire to reduce the nonproductive drill- 
ing time associated with pulling the rods 
out of the hole to change the bit. 

In the past, the commercial development 
of a retractable core bit system has been 
hampered by three broad categories of 
shortcomings: 

1. The cutting elements of the bits 
that were developed in the past retracted 
each time the inner core barrel assembly 
was pulled whether the bit was worn or 
not. With this type of system an ex- 
tremely high reliability factor is neces- 
sary to achieve an efficient operation. 



56 



1.^ 




L 



'«^P^«# 



FIGURE 1. - Schematic cross section of the retractable core bit drilling syster 



57 



2. The core bit, a consumable com- 
ponent, was conqjlicated, leading to 
relatively expensive machining, close 
tolerances, and inherent operating 
difficulties. 

3. The bit and associated retraction 
mechanisms were not rugged enough to 
endure the rigors of normal drilling 
operations. 



In order to increase core drilling 
productivity and advance the state-of- 
the-art in core drilling technology, 
the Bureau of Mines, through its con- 
tract research program and in-house capa- 
bilities, designed a down-hole replace- 
able core bit drilling system. The sys- 
tem has been fabricated and field 
tested. 



GENERAL DESCRIPTION 



This system functions as a conventional 
wireline core drilling system in all 
aspects until it becomes necessary or 
desirable to change the bit. After the 
inner core barrel is removed from the 
drill rods, the bit retraction tool is 
lowered through the rods by means of the 
wireline cable until it seats in the 
landing rig of the outer tube assembly. 
At the surface, a packer is applied to 
the top of the drill rods and fluid pres- 
sure is applied utilizing the available 
pumping system. This procedure locks the 
tool in place for the retraction opera- 
tion and assures that the tool is in the 
proper orientation. Raising the wireline 
cable, which remains attached to the 
tool, allows the retraction mechanism to 
lock into the bit, push it clear of the 
bit holder, and rotate it into position 
for passing through the drill rods 
(fig. 2) to the surface. The next step 
is to inspect and/or replace the bit and 
reverse the procedure as described above 
using the bit insertion tool. 

Both retraction-insertion tools are 
alike in appearance, size, and internal 
construction. The only difference is in 
the drive mechanism, which transforms the 
upward forces of the wireline cable into 
the combination of forces and movements 
needed to remove or replace the bit. In 
effect the internal drive mechanisms of 
the tools are reversed. 

System Conyonents and Operating 
Procedures 

The bit is a one-piece diamond set ele- 
ment position for drilling by two steel 
lugs and held by a locking device incor- 
porated within the bit holder of the 



outer tube assembly of the wireline core 
barrel. Its external configuration is 
similar to that of a wireline core bit 
except for omission of two portions that 
provide the interface with the driving 
lugs (fig. 3). 

The bit design allows for a multiplic- 
ity of external configurations, diamond 
settings, etc., just as with the wireline 
bit. The two basic differences between 
it and a conventional wireline bit are 
(1) in the manner of fastening to the 
core barrel outer tube assembly and (2) 
in the omission of part of the circumfer- 
ential surface in the area that abuts the 
driving lugs. 

The novel configuration of the bit 
allows it to be passed through the drill 
rods. The bit is removed from the drill- 
ing position and rotated in two planes, 
allowing its passage through the inside 
diameter of the drill rods. However, 
when the bit is in the drilling position, 
the hole and core are still cut in a con- 
ventional manner. 

So far as the drilling operation is 
concerned, the modified wireline core 
barrel (fig. 4) employed in the retract- 
able bit system operates exactly as con- 
ventional wireline core barrel. The 
basic modification is represented by the 
addition of a bit holder. This device is 
fitted on the lower end of the outer 
tube; when actuated by the retraction or 
insertion tool, it releases the bit for 
withdrawal or locks the new bit in place. 
It does not interfere with the normal 
function of the inner tube assembly (core 
barrel) . 



58 





A). Retractable bit after drilling. 



^'.-^ 



D). Retractable bit undergoing longitudinal and lateral rotations during retraction cycle. 





B). Retraction tool engaged, and ready to start retraction process. 



E). Retractable bit in final orientation at tfie end of the retraction cycle. 



m 




C.) Retraction tool pushing core bit clear of the bit holder. 



F). Retractable bit being pulled through the drills rods. 



FIGURE 2. - Bit retraction sequence starting from the drilling position (A) to the fully retracted 
position (F). 



59 




FIGURE 3. - Comparison of the retractable core bit (A) with a conventional wireline core bit (B). 



One minor modification to the wireline 
core barrel is the provision of a groove 
in the landing ring into which the in- 
serted tool is locked prior to its bit 
replacement or withdrawing operation. 
Another difference is a direct connection 
(no spring) between the inner tube and 
the spindle bearing. 

As a consequence of the retractable bit 
design, the diameter of the core cut in 



an N-size is reduced by 1/8 inch (3.2 mm) 
to 1.75 inch (44.45 mm). This is not 
considered a serious handicap. 

Since the retraction and insertion 
tools are identical in design and appear- 
ance except for one part, their operation 
will be described as one. The difference 
is that the direction of motion of the 
internal drive mechanism is reversed 
between the tools. 



60 



Locking coupling- 



Adapter coupling 



Landing ring- 



Outer tube 



& 



Reaming shell 



Head assembly 



Orienting pin and spring 



Solid spacer 




Inner tube 



Locking sleeve 



Core lifter case 



Detent spring 



Locking key 



FIGURE 4. - Schematic of the retractable bit system core barrel. (Unshaded parts are the 
same as wireline core barrel parts. Medium-shaded parts are changed in di- 
mension only. Dark-shaded parts are the new configurations.) 



61 



Simply stated, the retraction tool con- 
tains mechanisms that, after the tool 
is lowered and seats into the landing 
ring of the core barrel outer tube 
assembly: 

Lock the tool in place through 
fluid pressure making the tool ready 
for subsequent operations. 

Transform the pulling forces ap- 
plied by the wireline cable into a 
series of mechanical actions which 
grasp the bit; advance it past the 



end of the bit holder; rotate the bit 
through two perpendicular planes; and 
release the tool from the outer tube 
assembly, enabling the tool and bit 
to be withdrawn by the wireline 
cable. 

Externally the tools present the ap- 
pearance of a strong, rigid tube. Few 
moving parts are exposed. Internally, 
its function depends on the interaction 
of a number of sequential parts such 
as pins , cams , linkages , and drive 
mechanisms (figs. 5-7). 



-Spearhead 



Half-dog set screw - 



— Outer tube assembly 



■ Fluid bypass plug 



Camming grooves 



Tool lock balls 



FIGURE 5. - Schematic cross section of the up- 
per portion of the retraction tool. 




5ft (1.5m)— Extension 

sections are 
added here. 



Upper drive screw 



FIGURE 6. - Schematic cross section of the middle 
portion of the retraction tool. 



62 



Upper actuator 



'M 



Locking sleeve pin 



Upper drive nut 



180*' ROTATED VIEW 



Lower drive nut 



Lower drive screw 



Upper actuator 



Tool orienting grooves 



Longitudinal bit rotation pin 



Bit rotator 



Bit rotator' 



Vertical bit rotation pin 



, Vertical rotation linkage 




Ball plunger 



• Bit lock 



FIGURE 7. - Schematic cross section of the lower portion of the retraction tool 



63 



Both tools can accommodate the length 
of core barrel being used by adding ex- 
tension sections that fit between the 
upper and lower portions of the tool 
(that is, 5-, 10-, 15-foot core barrel). 
In practice, the procedures currently em- 
ployed in wireline core drilling apply to 
the replaceable bit system. Drill setup 
and collaring of the hole are accom- 
plished as before. 

By designing a replaceable bit with its 
associated retraction-insertion tools to 
function independently of the inner tube 
assembly, reliability of the system can 
be enhanced. This feature was an impor- 
tant design parameter. For example, the 
driller should be in a position to pull 
the core bit out of the hole for inspec- 
tion or replacement only when the deci- 
sion is made to do so. A simple example 
of this type of reliability is discussed 
below. 



pulling indicates that savings are 
realized where drilling conditions are 
greater than 1,000 feet (300 m) , under 
normal drilling conditions. However, 
time savings could be significant 
in holes less than 1,000 feet if bit 
life is poor (that is, less than 25 
ft/bit). These estimates are considered 
to be conservative, based on the pro- 
totype system. The development of a 
commercial system could reduce these 
figures. 

In addition to direct time savings 
through an increase in productivity, 
other indirect advantages are anticipated 
from a commercially available replaceable 
bit system: 

1 . Less driller fatigue. Changing 
the bit with the replaceable bit sys- 
tem requires very little physical 
effort. 



Assuming a 90% reliability of the sys- 
tem, and that the bit has to be retrieved 
each time the inner tube assembly is 
pulled the system could be expected to 
malfunction one out of 10 core runs, 
regardless of the footage drilled. How- 
ever, if the retractable bit were inde- 
pendent of the inner tube assembly 
(assuming the same reliability) , the 
driller would have to pull the rods out 
of the hole only once every 10 times that 
the bit was pulled for inspection as a 
result of drilling. This concept sounds 
elementary, yet a historical look at past 
retractable bit designs and prototypes 
shows that almost all of the bits func- 
tioned as an integral part of the inner 
tube assembly, which contributed to poor 
reliability and lack of success. 

Benefits of the Replaceable Bit System 



2. Lower fuel consumption. Pulling 
rods requires engine operations that re- 
sult in high fuel consumption. 

3. Improved safety. Pulling rods is 
the major cause of injuries to drillers 
and helpers. 

4. Improved drilling efficiency. 
Drillers will tend to change bits to max- 
imize drilling efficiency either because 
of formation changes or as the bits be- 
come dull. 

5. Less time required to complete 
hole. Savings in drilling costs due 
to faster drilling speeds are only part 
of the savings. Savings associated 
with maintaining support personnel and 
facility onsite could also be 
significant. 



A comparison of the time savings of a 
replaceable bit versus conventional rod 



64 



CONCLUSIONS AND CURRENT STATUS 

The retractable bit system has under- as well as engineering evaluation studies 

gone field testing under a variety of to determine the overall commercial feas- 

conditions and rock types. The results ibility of the system. During 1981, 

of the initial field program are very field tests will be conducted on contract 

encouraging and indicate that the re- drill sites under production conditions, 

tractable bit system is feasible and Domestic and foreign patents on the sys- 

reliable. An extensive field program by tem have been applied for by the Longyear 

the Longyear Co. is currently underway, Co. 



I 



65 



STRUCTURAL DESIGN FOR DEEP SHAFTS IN HARD ROCKS 
By Michael J. Beus "• and Samuel S. M. Chan2 



INTRODUCTION 



Shaft design has traditionally meant 
specifications of shaft dimensions, 
hoisting plant and capacity, skip, guide, 
and headframe design, sinking method, and 
location. In recent years there has 
developed a need for shafts of more 
sophisticated design In both the estab- 
lished mining areas and the newer dis- 
tricts. Deeper and lower grade ore 
bodies require faster sinking methods, 
higher tonnage, lower maintenance, and 
Increased ventilation. 

Deep mine shafts frequently experience 
heavy ground pressure, and failure of the 
rock and/or supporting structures occur 
both during sinking and throughout the 
life of the shaft. This requires almost 
continuous shaft repair and maintenance, 
exposure to hazardous working conditions, 
and excessive operating costs. 

The Bureau of Mines is conducting 
research to develop guidelines for shaft 
design to improve the structural behavior 
in deep, vein-type metal mines. The 
primary objectives are — 

1. Define the nature, magnitude, and 
direction of the stresses acting around 
proposed shaft openings. 



2. Determine the structural sensitiv- 
ity of various shaft designs to changes 
in applied load, shape, orientation, and 
support. 

3. Conduct prototype and full-scale 
field studies. 

4. Establish design criteria. 

Initial studies concentrated on in situ 
stress measurement and determination of 
physical properties. These basic data 
were developed into predictive equations, 
and an appropriate range of input values 
for finite-element method analysis (FEM) 
was established. By comparing strength 
of the rock and supporting materials, 
the stability of the shaft may be pre- 
dicted. Various parameters are consid- 
ered including shape, orientation, 
stress ratio, geologic discontinuities, 
support system, and time to support. 
An example shows how a hypothetical 
shaft might be designed for maximum 
stability. Prototype field tests com- 
pared deformation of circular and rec- 
tangular one-half scale test shafts, 
taking into account geologic and con- 
struction variables. 



SITE DESCRIPTION 



The Coeur d'Alene mining district in 
northern Idaho is being used as a test 
area. It is typical of a deep-vein min- 
ing area experiencing shaft stability 
problems, and many shaft test sites are 
available. The district is situated in 
the Coeur d'Alene Mountains in northern 

^Mining engineer, Spokane Research Cen- 
ter, Bureau of Mines, Spokane, Wash. 

^Professor of Mining Engineering, Uni- 
versity of Idaho, Moscow, Idaho; WAE at 
Spokane Research Center, Bureau of Mines, 
Spokane, Wash. 



Idaho, It is mountainous with peaks 
ranging from 6,000 to 7,000 ft and a 
regional relief of 3,000 to 4,000 ft. 
The main rock in the district is Pre- 
cambrian quartzlte and argilllte with a 
maximum thickness of 28,000 ft. The main 
structural feature is the Osburn Fault, 
which strikes west-northwest and has 
extensive displacement. The major fold 
is the Big Creek anticline, south of the 
Osburn Fault. Most major mines in the 
district are located near the Osburn 
Fault or its branches. The min- 
eral deposits occur as steeply dipping. 



66 



quartz-siderlte veins containing silver- 
bearing tetrahedrite, galena, and sphal- 
erite. Most of the veins are paral- 
lel to the bedding or cut the bedding at 
small angles. 

The main mining method is horizontal 
cut-and-fill s toping using conven- 
tional drill, blast, and mucking cycles. 
Hoisting and rail haulage are the major 



means for personnel and materials trans- 
portation. Rock bolting, timbering, and 
hydraulic sandfilling are the main types 
of ground support. Shafts in the dis- 
trict extend to more than 3,000 ft below 
sea level and more than 8,000 ft below 
the ground surface. Figure 1 shows an 
idealized cross section of major mines 
and shafts in the district with respect 
to sea level and adjacent topography. 



STRESS DETERMINATION 



Determination of the magnitude and 
direction of in situ stresses is essen- 
tial to reliable design studies. Mea- 
surement has been made at various sites 
in the Coeur d'Alene District ranging in 
depth from 1,200 to 7,000 ft below the 
ground surface. Table 1 shows the prin- 
cipal stress ratios in decreasing order, 
with respect to depth. These data show 
that the stress conditions are not uni- 
form and the principal stress ratio is 
not a function of depth, because of tec- 
tonic activity influencing the regional 
stress pattern. 



TABLE 1 . - Principal stress ratios 
determined in the Coeur 
d'Alene mining district 



Mine 


Principal 

stress ratio 

(01/03) 


Depth 
of test 
site, ft 




3.18 
3.11 
2.80 
2.65 
1.85 
1.78 
1.25 


4 800 


Lucky Friday 

Silver Summit.... 

Caladay 

Galena 

Star 


4,250 
5,500 
1,220 
4,000 
7 340 


Crescent 


5,300 



The in situ stress information is fur- 
ther reduced to vertical and horizontal 
components in table 2, shown in order of 
increasing depth. The ratio between 
horizontal and vertical stress and 
maximum and minimum horizontal 
stresses is also shown. Failure pat- 
terns in vertical raise bores in the 
district illustrate the strong biaxial 
stress condition. Figure 2 is an example 
of shear failure in a 5-ft-diam raise 
bore, resulting from a strongly direc- 
tional stress acting 90° from the failed 
zones. 



TABLE 2. - Vertical and horizontal 
stresses, Coeur d'Alene 
mining district 





Over- 


Vertical 


^h, / 


^hi / 


Test 


burden, 


stress 


/ 


y 


site 


ft 


Ov, psi 


/a. 


/^h2 


Caladay.. 


1,200 


1,450 


0.88 


1.56 


Galena... 


4,000 


5,500 


2.37 


1.36 


Lucky 










Friday. . 


4,250 


4,770 


2.00 


1.42 


Sunshine . 


4,800 


7,420 


.97 


1.81 


Crescent. 


5,300 


6,300 


1.24 


1.25 


Silver 










Summi t . . 


5,500 


7,870 


1.87 


2.73 


Star 


7,340 


7,280 


1.43 


1.52 



By statistical analysis of the data 
from table 2, linear relationships may be 
derived enabling prediction of vertical 
and horizontal stresses at any depth. 
The expression for vertical stress (Oy) 
becomes 



a„ = 435 + 0.952 h. 



(1) 



and that for maximum horizontal stress 
(a. ) is 



710 + 1.491 h. 



(2) 



All of the data analyzed through 1980 
show that — 

1. The vertical stress is comparable 
to what might be expected from a gravity- 
loaded mass. 

2. The horizontal stresses are greater 
than the vertical. 

3. The horizontal stress ratio ranges 
from 1:1 to almost 3:1. 



67 




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CM CO ^IO<0 

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68 



Va«*:^'i: :^^i = -< ■^i5-^.^:f«^v?,S. 




FIGURE 2. - Shear failure in a 5-ft-diam raise bore. 



Consideration of the in situ stress field 
in the overall shaft design procedure is 
therefore of prime importance, partic- 
ularly in deep mining. The increasing 
stress levels could cause stress concen- 
trations around the opening exceeding the 
rock strength. 



1 . The rocks are 
deform elastically. 



hard, brittle, and 



2. The strength of the rocks and rock 
masses is variable depending upon the 
mineral content, fracturing, and inter- 
layered argillite. 



Physical properties of rocks from the 
various mines of the Coeur d'Alene mining 
district have also been determined, both 
in situ and in the laboratory. Summariz- 
ing these tests on competent quartzite 
specimens, it has been found that — 



3. The physical properties vary, 
ticularly with lateral confinement. 



par- 



4. The rocks creep at very small rates 
of deformation. 



FINITE-ELEMENT ANALYSIS 



A variety of FEM models has been 
developed to illustrate stresses around 
different shaft shapes and support 



systems, under nonuniform loading condi- 
tions, and for different types of rock 
behavior. Circular, rectangular, and 



69 



elliptical cross sections are the three 
major types of shaft shapes studied. 
Unsupported and supported shafts with 
timber, steel, and concrete lining were 
simulated. The size of the support 
structure was also varied. 

Initial analyses consisted of paramet- 
ric studies to establish the sensitivity 



of "variable" and "fixed" conditions. 
Fixed conditions include applied loads, 
physical properties, and geologic 
features; design variables are shape, 
axis length, orientation, support, 
and face distance. Table 3 lists physi- 
cal properties used for the parametric 
s tudy . 



TABLE 3. - Design input parameters for finite-element analysis 





Modulus of 


Poisson's 


Bulk 


Shear 


Materials 


elasticity 


ratio, y 


modulus 


modulus 




(E), psi 




(B), psi 


(G), psi 


Hard quartzite 


10,000,000 


0.25 


6,666,700 


4,000,000 


Arglllite 


1,000,000 


.25 


666,670 


333,340 


3,000-psi concrete 


3,140,000 


.15 


1,495,240 


1,365,220 


9,000-psi concrete 


5,407,500 


.15 


2,575,000 


2,351,000 




30,000,000 
1,900,000 


.3 
.25 


25,000,000 
1,260,000 


11,538,461 


Douglas-fir timber 


760,000 


Sand 


2,000 


.15 


952 


870 



Support loads occur as a result of 
changes in stress state and rock proper- 
ties and deformation of the shaft wall 
rock associated with face advance. 
Either a horizontal or a vertical cross 
section is possible with a two- 
dimensional analysis. A vertical plane 
through the shaft centerline permits 
analysis of the effects of face advance. 
A horizontal section permits study of the 
shape of the cross section, orientation, 
and support, assuming the support is 
installed concurrently with excavation. 
Obviously, this results in considerably 
higher stresses than would actually be 
encountered; however, for comparative 
purposes, this approach is valid. 

A two-dimensional, 14-ft-diam, circular 
shaft model illustrates the simple case 
of variable applied stress in lined and 
unlined conditions. Figure 3 shows the 
influence of the applied stress ratio and 
the effect of a 2-ft-thick concrete 
liner. Design procedures for circular 
shafts in an elastic medium are fairly 
straightforward, and further analysis is 
not presented here. 

The remainder of this paper emphasizes 
rectangular shafts, as this is where de- 
sign procedures and analysis are lacking. 
The primary structural consideration for 



rectangular shafts is orientation, with 
respect both to the prevailing stress 
field and to discontinuities in the rock 



The most stable orientation for a 
rectangular-shaped opening in a biaxial 
(2:1 or 3:1) stress field is where the 
long axis is parallel to the direction of 
the largest stress. The least stable 
situation is unidirectional loading, re- 
gardless of orientation, or with major 
stress normal to the longwall of the 
shaft. Loads up to 9,000 psi have been 
applied to shaft sections at various 
ratios and orientation. The preference 
for orientation remains the same, except 
that, for increasing stress levels, 
orientation becomes less a factor as the 
yield zone increases around the opening. 
The general trend is towards an increase 
in yield as the major axis of the shaft 
is rotated from parallel to perpendicular 
to the major stress. 

Geologic discontinuities such as bed- 
ding planes, joints, and/or faults also 
have a significant effect on shaft sta- 
bility. A condition of interbedded 
quartzite and argillite, striking paral- 
lel with the long axis of a rectangular 
shaft, is shown in figure 4. The tangen- 
tial stress and axial displacement are 



70 



20 



16 



lO 


1? 


O 




^ 




« 




w 




LU 


8 


q: 





(f) 



KEY 

Applied stress (P,/Py),psi 6,000/6,000 6,000/2.000 

Unsupported ■ > 

2-ft concrete lining o o 



90* 




67.5** 45" 22.5* 

ELEMENT POSITION, degrees from horizontal 



FIGURE 3. - Tangential stress distribution in rocks around a 14-ft-ID supported and unsupported 
shaft. 



^ 



10 



Quartzite 



>Arglllite 



Mm^mmmmmmm 



FIGURE 4. - Model of interbedded quartzite and argillite striking parallel with the long axis of the 
shaft. 



71 



considerably greater for the interbedded 
quartzite and arglllite than for quartz- 
ite alone. The most extensive displace- 
ment occurs where geologic discontinu- 
ities strike parallel to the major shaft 
axis. 

The best shaft orientation is to 
orient the long axis of the shaft 
parallel with the major stress and 
perpendicular to bedding as predominant 
joint patterns in the horizontal plane. 
Obviously, this can be accomplished only 
if the strike of the bedding and joints 
is normal to the major stress. 
Ultimately, it is envisioned that the 
best orientation will be a compromise 
between consideration of the prevail- 
ing stress field and of geologic 
discontinuities . 



Additional studies evaluated support 
systems commonly installed in shafts. 
Timber, steel, or concrete act as a 
structural framework to mount the shaft 
conveyances and provide protection from 
loose blocks of falling ground. There is 
little change in stress concentration 
factor in the rock as a result of 
installation of these support systems, as 
illustrated by figure 5. The steel and 
timber supports result in local high 
stress zones, particularly at blocking 
points. The concrete liner results in a 
more uniform stress distribution around 
the opening. For timber and steel sup- 
ports, the highest stress occurs at the 
dividers. The steel set with timber 
blocking is very effective in reducing 
stress concentration radial displacement 
in the rock. 



2 5 



i5 

Z 3 

O 



UJ 

o 

z 
o 
o 

(O 
CO 



2 - 



I - 



KEY 

-o— Unsupported 

—<t— Concrete 

-o— Timber 

-o- Steel 



j 3,000 psi 

lO'l 1- 

U 20' J 




CO 



ELEMENT POSITION 

FIGURE 5. - Comparison of supported and unsupported rectangular shafts. 



72 



DESIGN ILLUSTRATION 



Finite-element analysis may be applied 
to investigate a wide range of shaft 
design parameters, and it is still the 
most economical approach to structural 
analysis. However, it would be impracti- 
cal to analyze all the various design 
combinations. To permit design decisions 
based on structural performance for a 
given set of input conditions, the term 
"critical depth" is suggested. This term 
has been previously used by Trollope to 
define the maximum depth that a circular 
opening would be stable in elastic ground 
under uniform loading. In the context 
used here, it is the minimum depth at 
which material first yields, according to 
the Mohr-Coulomb yield criteria. For 
deep shafts, this concept is especially 
appropriate, since the ultimate depth of 
mining is undefined and the shaft design 
permitting the deepest penetration is 
probably the most appropriate. In con- 
trast, if the vertical extent of the ore 
body, and thus the maximum depth of the 
shaft, is defined, such as in a massive 
deposit, any number of designs would be 
suitable. 

The following example illustrates the 
design procedure. The location of the 
shaft has been predetermined, the shaft 
station has been excavated, and the 
essential physical properties and in situ 
ground stress have been measured. The 
project is situated in mountainous ter- 
rain, and the shaft station is under 
1,220 ft of quartzite, connected to the 
surface through a 5,000-ft adit. 

The physical properties of the quartz- 
ite are as follows: 

unconfined compressive strength, C^ 
= 18,275 psi, 

tensile strength, T^, 1,754 psi, 

modulus of elasticity, E = 9.57 
X 106 psi^ 



Poisson's ratio, y = 0.205, 
internal friction angle, (^ = 52.2°, 
cohesion, C = 2,836 psi, and 
failure plane angle, 9 = 60°. 
The in situ stress is 



and 



1,450 psi 



Oh^ = 820 psi. 



level I 
ontal j 



From equations 1 and 2, stresses at any 
depth are estimated. The vertical stress 
at the 3000 level is 4,452 psi, and the 
maximum horizontal stress is 7,002 psi. 
The ratio of a^ /o^ is 1.6, so the mini- 
mum horizontal stress at the 3000 
is 4,489, assuming that the horizon 
stress ratio does not reorient with 
depth. These data provide the basic FEM 
input. 

Analysis of supported and unsupported 
14-ft-diam circular shafts show plastic 
yield beginning between the 1,400- and 
1,500-ft levels at a critical depth of 
2,700 ft. The critical depth for a 
"favorably oriented" 10- by 20-ft, unsup- 
ported, rectangular shaft is 4,770 ft. 
Where the shaft is supported by timber 
sets, it can go as deep as 4,920 ft. 
The critical depth of the same shaft 
when supported by steel sets is 
6,820 ft. 

This is only a preliminary approach and 
these results must be used with caution 
because limitations in the FEM code and 
computer capability prevent simulation of 
unequal loading and shaft advance at the 
same time. The procedure is being fur- 
ther developed, and future work will 
incorporate construction variables, un- 
equal loading, and nonelastic material 
properties. 



ONE-HALF- SCALE TEST SHAFTS 



73 



Mathematical analysis and model studies 
are invaluable for assessing design 
alternatives and the effect of various 
input parameters. FEM studies provide a 
comparative basis of design parameters. 
However, the best way to accurately pre- 
dict the behavior of a full-size opening 
is by construction and monitoring of a 
scaled prototype. The actual deforma- 
tional behavior of circular and rectangu- 
lar test shafts has been compared by 
in situ testing. The effect of the con- 
struction method and geologic defects, 
difficult to assess by mathematical or 
physical modeling, were determined. The 
experiment simulated, as closely as 



possible, 
shaft. 



the excavation of an actual 



Circular and rectangular shaft openings 
of similar cross-sectional area were 
excavated at the 1,200-ft depth in an 
existing shaft station. An 8-ft-diam 
circular shaft and a 5- by 10-ft rectan- 
gular shaft were sunk to a depth of about 
27 ft (fig. 6). A full-size shaft at the 
same orientation, with the long axis nor- 
mal to the strike of the bedding, was 
proposed at the site of the rectangular 
test shaft. Figure 7 is a schematic of 
the proposed test shaft excavation 
sequence for both shafts. 




ent site: 

at N 87° E 
at S 3° E 



FIGURE 6. - Plan view of proposed test shafts, hoist room, and connecting drifts in the Caladay 
Project, 



74 




FIGURE 7- = Schematic of the station and test shafts. 



Epoxy-filled 
glass jar 



Deformation sensing plug 



Threaded stock 




Support reference pipe 



Adjustable 
mounting head 



FIGURE 8. - Components of the TSR measurement system. 



The instrument used in the test was 
developed and patented by the Bureau in 
1971. It is called a tunnel stress 
relaxation (TSR) gage and is designed 
to measure the initial and long-term 
deformation around large underground 
excavations. It consists of the TSR 



probe, a deformation sensor plug, an 
adjustable mounting head, and an anchored 
support reference pipe (fig. 8). The 
sensing plug is epoxied into the end of a 
borehole parallel with the mine opening 
and fits the ball end of the transducer. 
Movement of the plug causes deflection of 



75 



the probe. Figure 9 shows the Instru- 
ments installed around the completed 
circular shaft. The data acquisition 
station was located in an instrumenta- 
tion room downdrift from the shafts. 
Blasting was controlled from this room, 
and detonation was initiated with the 
start of the data acquisition cycle to 
obtain maximum data. 

After initializing the instruments, 
the round was detonated, readings 
allowed to stabilize, and the process 



repeated. Shaft muck was removed with 
a backhoe and slusher-clam bucket and 
loaded into a train loader for transport 
to the surface. Final depth of the cir- 
cular shaft after eight rounds, over a 
period of about 2 weeks, was 24 ft, 
roughly two opening diameters (16 ft) 
beyond the TSR sensing plane (fig. 9). 
The rectangular shaft was excavated with 
seven consecutive rounds to a final 
approximate depth of 26 ft, roughly 17 ft 
beyond the TSR sensing plane (fig. 10). 




FIGURE 9. - The completed circular test shaft and the installed instrumentation. 



76 





FIGURE 10 = The installed instrumentation along the long axis of the rectangular shaft. 



77 



Each blast developed a characteristic 
pattern, or trend, during the blast. 
Rock response could be classified as 
either elastic or inelastic. A rela- 
tively slow and small deformation occur- 
red for about 1 hr following each blast 
at a slowly decreasing rate. Figures 11 
and 12 show the displacement traces for 
the circular and rectangular shafts, 
respectively. 



diameters (about 17 ft) beyond the TSR 
sensing plane. A 4-month monitoring 
period following completion of the open- 
ings showed no noticeable creep charac- 
teristics. The smallest displacements 
were measured at the 0° and 45° positions 
in the circular shaft (parallel and 45° 
to the bedding) and between the center 
and the corner of the long axis of the 
rectangular shaft. 



Considerable difficulty was experienced 
with excessive water and the inherent 
complications resulting from the use of 
explosives and their effect on the 
instrumentation. Despite these diffi- 
culties, valid data were obtained, and 
some significant data trends were noted. 
Maximum displacement occurred normal to 
the major geologic planes of weakness 
(bedding planes), and in the case of the 
rectangular shaft, was extreme to the 
point of endwall failure. Elastic insta- 
bility was evident as beam bending at the 
midpoint of the long axis of the rectan- 
gular shaft. Displacements generally 
ceased as the face advanced two-opening 

ROUND NO 



Comparing the actual data record 
for the circular and rectangular test 
shaft, the following conclusions regard- 
ing their relative stability are 
appropriate. 

1. The circular shaft was less sensi- 
tive to the excavation process, as shown 
by the subdued nature of the deformation 
record. 

ROUND NO. 



09 


1 1 1 1 


1 


06 


Afl 




03 


- h / K^^^-^ 


^--v-^'- ^r 





l>-l/ f^ ^^^^ 


003 


1 1 1 1 1 




FIGURE lit - Deformation around the circular test 
shaft measured with the TSR's as a 
function of shaft depth. 



FIGURE 12. - Deformation around the rectangular 
test shaft measured with the TSR's 
as a function of shaft depth. 



78 



2. The magnitude of displacement was 
generally smaller and more uniform around 
the circular shaft and less affected by 
geologic discontinuities. 



3. The rectangular shape is more sus- 
ceptible to the effects of elastic 
instability; that is, beam bending and 
buckling, and geologic discontinuities. 



SUMMARY 



This research centers around the con- 
troversy of circular versus rectangular 
shaft shapes. Stress and physical prop- 
erty measurements have shovm that the 
horizontal in situ stress around a shaft 
can be several times larger than the 
vertical stress. The horizontal stresses 
are unequal, with up to a 3:1 ratio 
between major and minor stresses. These 
data justify detailed investigations of 
various shaft designs to develop long- 
term stability of the opening. 

With the FEM technique, the structural 
implications of various design and con- 
struction variables are investigated 
using actual field data as input. Many 
design options are available that satisfy 
stress concentration criteria for a given 
set of input conditions. To illustrate a 
design procedure that minimizes the 
design variables, the term "critical 
depth" is used. This concept is based on 
the premise that the single "best" shaft 
design is one that permits the deepest 
penetration of the deposit, an approach 



unique to a steeply dipping, vein-type 
ore body. Selection is based on struc- 
tural performance in plastic ground con- 
ditions with unequal horizontal loads. 

It is difficult to mathematically model 
actual construction practices and geo- 
logic features. Field testing of scaled 
test shafts has allowed investigation of 
these parameters for circular and rectan- 
gular shaft shapes. 

The FEM approach is being improved to 
more realistically model actual construc- 
tion practices and rock structure. A new 
computer at the Bureau's Spokane Research 
Center and modified FEM codes will permit 
improved structural analysis. In addi- 
tion, full-sized circular and rectangular 
shafts are now being instrumented to 
fully describe the structural behavior of 
the shaft and support system. An 
improved shaft design method should 
emerge from these studies as further data 
are obtained and compared with currently 
used criteria. 



79 



BOREHOLE DEVIATION CONTROL 
By E. H. Skinner 1 and N. P. Callas2 



INTRODUCTION 



This paper presents the results of con- 
tract research for the past 2 years at 
the Spokane Research Center (2^). Two 
areas of borehole research have been 
undertaken: (1) the mathematic analysis 
for the preferred method of calculating 
borehole surveys, and (2) the 2-D analy- 
sis of drill-string mechanics. The text 
of this paper will briefly summarize the 
results of each of these areas. The 
Hewlett-Packard (HP) and Texas Instru- 
ments (TI) computer programs for the 
borehole survey calculation method are 
available upon request from the Spokane 
Research Center. An appendix briefly 
covers the elementary theory of drill- 
string mechanics. 

It was from the review of the litera- 
ture that our research plan developed. 
It was noted that many papers addressed 
problems of borehole surveys. It is 
well acknowledged that mining applica- 
tions of drilling require much more 
precision in drilling and surveying 
than most other drilling, such as oil 
field-type drilling. Another problem 
noted in the literature was the drilling 
of straight holes to specific underground 
targets and the problems of drill-pipe 
twistoffs, particularly at the bit end 
and at stabilizers that were placed near 
the drill bit. We believe that we have 
now addressed each of these problems to 
the limits of present technology (5). 

Straight-hole drilling may use inten- 
tionally deviated drill holes to help 
achieve optimal drill-hole usage and, 
particularly, will extend the capability 
of surface drilling equipment. In many 
cases, carefully engineered directional 
drilling offers the only method for 

"Mining engineer, Spokane Research 
Center, Bureau of Mines, Spokane, Wash. 

^Department of Mathematics, Colorado 
School of Mines, Golden, Colo. 



reaching desired target locations. How- 
ever, as the borehole inclination angle 
increases, peculiar drilling problems 
begin to develop that severely limit 
the horizontal extent of directional 
drilling. Among these problems are (1) a 
general tendency to lose the ability to 
control bit weight and provide bit guid- 
ance; (2) optimum rate of penetration is 
lost and, therefore, economic rig use is 
lost; (3) the ability to clean the drill 
hole is decreased and maintaining effec- 
tive drilling hydraulics becomes a prob- 
lem; (4) drill pipe sticking in the bore- 
hole; and (5) general difficulties in 
downhole movement of drill pipe and cas- 
ing, and in wireline tools. The advan- 
tages in being able to drill to great 
horizontal distances at controlled angles 
is that the effective drilling area may 
be increased by as much as 10 times. 
Certainly, the technique could find 
application in a number of special drill- 
ing applications. 

A list of some of the possible mining 
uses for straight-hole drilling, con- 
trolled directional drilling, as well as 
general mine drilling applications, is as 
follows: 

1. Geologic data collection. 

2. Ore-body definition. 

3. Ore reserve analysis. 

4. Surface environmental 
considerations . 

5. Hazard detection and/or evasion 
drilling. 

6. Health and safety emergency exits. 

7. Water inflow and shutoff control 
methods. 



80 



8. Closely controlled vertical and 
inclined shaft predrilling. 

9. Ore pass development. 

10. Drilling for ventilation openings. 

11. Manway raises and stope raise 
development. 

12. Large-scale blasthole mining 
methods. 

13. Boundary drilling for bulk mining 
methods . 



may not allow total development of this 
concept at every mine. 

However, some of the advantages are the ^ 
following: 

1. Eliminates one 200-ft level devel- 
opment with time and cost savings for all 
development expense. 

2. Removes one less rock-burst-prone 
sill mining sequence. 

3. Recovers all ore in one 40-ft sill 
interval now lost with the 200-ft level. 



14. Drilling for induced ore caving. 

15. Special drill holes for running 
utilities. 

16. Special drill holes for waste 
backfill. 

17. Special drill holes for hydraulic 
pumping of ore. 

18. Special drill holes for firefight- 
ing during a mine disaster. 

19. Multiple drill holes from one cen- 
tral drilling location. 

20. Drilling for the promising new 
method of solution mining. 

Each of these special drill-hole pro- 
grams will require different engineer- 
ing criteria. To illustrate one special 
application of straight-hole drilling 
with innovative mine development con- 
sequences is the following ore pass 
development exclusively through raise- 
boreholes using straight-hole pilot- 
hole drilling practices. This example 
(fig. 1) proposes use of 400-ft level 
development instead of the customary 200- 
ft level development; thus, we eliminate 
one level from the present system, which 
will allow 100% recovery of ore from 
one sill and vast savings in time and 
cost for level development and cutting 
shaft stations. Of course, ore-body 
geometry and many other considerations 



4. Enables time and cost savings in 
station cutting and drift development for 
one level towards the vein. Also elim- 
inates a tremendous amount of waste muck 
removal from the extra 200-ft level. 

5. Consolidates hoisting from 400-ft 
levels rather than from 200-ft levels, 
which may be better programed for 
automatic hoisting control from fewer 
levels. 

6. The circular ore pass with an in- 
clination of 50° to 80°, and a diameter 
of 5 ft or more is a preferred opening 
for an ore pass. 

7. Savings in level timber consumption 
and other supplies for the 200-ft level 
development. 

8. Raise-bore cuttings are ideal for 
backfill use and roadway surfaces. 

The disadvantages are — 

1. Requires accurate pilot hole 
drilling. 

2. May require a lined muck pass 
design. 

3. Raise-bore holes totaling more than 
1,000 ft may be seen as time consuming to 
complete. 

4. May experience muck hangups in long 
ore passes. 



81 




^ 



\6 



^^'^ ^' . Ke^4^ 






1 U©^ 





. 1 



Rafse-bore ore passes 



FIGURE 1. - Proposed ore pass mine development using raise=bore hok 



82 



5. May experience rock mechanics prob- 7. Better mine planning required, 
lems in highly stressed ore bodies. 

8. More engineering effort expended in 

6. More accurate knowledge of ore-body loading pockets at 400-ft intervals, 
geometry required. 

DRILL SURVEY CALCULATION METHODS 



Instruments of various kinds for the 
surveying of drill holes have been used 
since the beginning of modern drilling 
and today are the subject of massive R&D 
efforts by industry brought on by the en- 
ergy crisis. An associated subject area, 
which was apparently largely overlooked 
until the early 1960's, is the "best" 
mathematical presentation of borehole 
survey results. A profusion of various 
methods has been developed since the 
1960's, many of which are very similar, 
and indeed the work presented herein con- 
siders only seven methods. The most 
accurate method has been documented as a 
Fortran computer program and implements a 
vector approach to the data. The vector 
approach is a unique contribution from 
this research and offers calculations 
free of the usual anxiety in arithmetic 
calculations based on azimuth angles. In 
addition, a handheld computer program was 
developed that implements the same vector 
approach and offers all the advantages of 
computations of the larger computer pro- 
gram. These are available to industry 
for either the HP-67/97 or the TI-58/59. 
It is our purpose in presenting these 
programs, as well as the subsequent dis- 
cussion on drill stabilizers, to present 
the means to industry to perform these 
calculations right at jobsite with a 
handheld calculator. 

Drill-hole surveying methods have come 
to be known as two-point methods because 
successive departures are computed by 
using the directional survey data from 
adjacent stations to compute the next 
incremental horizontal and vertical posi- 
tions from the known previous station 
point. At each station, then, there are 
the usual three measured quantities of 
elevation, inclination, and azimuth, 
which are the customary directional sur- 
vey data associated with that station. 
Nothing has been changed from the previ- 
ous collection of survey data. 



The methods (4_) that have been analyzed 
are the following: 

1. Circular arc. 

2. Minimum curvature. 

3. Average angle. 

4. Balanced tangential. 

5. Radius of curvature. 

6. Walstrom, model 1. 

7. Walstrom, model 2. 

Unit vectors are taken as directions 
along the axis of the well bore at adja- 
cent survey stations. The angle between 
the vectors is the dogleg severity and is 
calculated by applying the arc cosine 
function to the dot product of the unit 
vectors. Hence, the vector departure 
along a uniquely defined circular arc 
falls in the chordal direction of the 
unit vector. The chord length is also 
calculated from the difference in the 
measured depths. The vector increment of 
departure between two stations can be 
given by a vector formula; and the final 
form, so obtained, is exactly the same 
for the circular arc and the minimum cur- 
vature method. This verified that these 
two models are exactly identical and 
would give identical departure results. 

In the sequel that developed the other 
models, it is convenient to examine the 
normalized circular departure vector. 
This vector will be compared with the 
normalized departure vectors among the 
other five distinctive methods. In this 
manner, upper bounds for the normalized 
differences of departure results between 
pairs of models were developed. Absolute 
differences can then be calculated from 
these models by merely taking the maximum 



83 



of the upper bounds and multiplying the 
normalized differences by the measured 
length between stations. The mathemati- 
cal development of these normalized 
departure formulas has been presented by 
Callas (4_, 9^) and will not be reviewed 
herein except to note that the upper 
bounds between differences were obtained 
by expanding the X, Y, and Z terms as 
two-dimensional Taylor sine and cosine 
function series in terms of the differ- 
ences of inclination and azimuth. Inci- 
dentally, the first two terms of these 
expansions can be used as an approxima- 
tion formula for the dogleg severity 
angle with appropriate change of units to 
degrees per 100 ft. Formulas for the 
normalized departures X, Y, and Z, in 
terms of these series expansion differ- 
ences of inclination and azimuth, are 
given in tables 1, 2, and 3, with coeffi- 
cients of each expansion tabulated for 
each of the six distinct models. Note 
that the 0th and the 1st order terms for 



all of the methods are identical, thus 
verifying the approximate equivalency of 
each two-point method. Different coeffi- 
cients arise when higher order terms are 
considered. For small unit values of 
inclination and azimuth, these higher 
order terms are small in comparison with 
the lower order terms. Hence, the dif- 
ferences are also relatively small. 
Therefore, the comparison between pairs 
of models reveals that they are exactly 
equivalent up to, but not including, 
second-order terms when their formulas 
are considered as series expansion of the 
differences of successive inclination and 
azimuth angles. These results confirm 
the long-held opinion among many direc- 
tional surveyors that there is little 
difference between departures computed by 
the various two-point methods. Recall, 
too, that the results also present a sim- 
ply applied approximation for computation 
of dogleg severity angle. 



TABLE 1 . - Normalized departures for X 



Circular 



Average 
angle 



Balanced 
tangential 



Radius of 
curvature 



Model 4 Model 5 



ao.. 
aio- 
aoi< 
320. 
an. 
ao2' 
bo2' 
asO" 
ai2- 
bi2' 
^21- 
ao3' 
bo3' 



1 
1/2 
1/2 
•1/6 
1/2 
•1/4 
1/12 
•1/24 
•1/4 
1/8 
•5/24 
•1/12 
1/24 



1 

1/2 

1/2 

-1/8 

1/4 

-1/8 


-1/48 
-1/16 



-1/16 

-1/48 





1 
1/2 
1/2 
•1/4 
1/2 
•1/4 


•1/12 
•1/4 


■1/4 
•1/12 





1 
1/2 
1/2 
■1/6 
1/4 
■1/6 


■1/24 
•1/12 


■1/12 
•1/24 





1 

1/2 
1/2 
•1/6 
1/3 
■1/6 


•1/24 
•1/8 


•1/8 
•1/24 





1 

1/2 

1/2 

-1/4 

1/4 

-1/4 


-1/12 
-1/8 



-1/8 

-1/12 





X = 



sin I sin A + ai n cos I sin A dl -f am sin I cos A dA 



■10 



01 



+ a2o sin I sin A dl^ +3^1 cos I cos A dl dA + ao2 sin I sin A dA' 

+ bQ2 sin^ I sin A dA^ + 339 cos I sin A dl^ 

+ a2i sin I cos A dl^ dA -f ai2 cos I sin A dl dA^ 

+ bi2 cos I sin^ I sin A dl dA^ "•" ^03 sin I cos A dA^ 



+ b 



3 sin^ I cos A dA^ + Oi 



84 



TABLE 2, 



Normalized departures for Y 



^0 


^10 


-^01 


^20 


-an 


^02 

^02 


^30 


^12 

^12 


-^21 


-^03 

-b03 



Circular 
arc 



1 
1/2 
■1/2 
■1/6 
■1/2 
■1/4 
1/12 
■1/24 
■1/4 
1/8 
5/24 
1/12 
-1/24 



Average 
angle 



1 
1/2 
-1/2 
-1/8 
-1/4 
-1/8 



-1/48 

-1/16 



1/16 

1/48 





Balanced 
tangential 



1 
1/2 
-1/2 
-1/4 
-1/2 
-1/4 



-1/12 

-1/4 



1/4 

1/12 





Radius of 
curvature 



1 

1/2 
-1/2 
-1/6 
-1/4 
-1/6 



-1/24 

-1/12 



1/12 

1/24 





Model 4 Model 5 



1 

1/2 
•1/2 
•1/6 
•1/3 
•1/6 


■1/24 
•1/8 



1/8 
1/24 





1 
1/2 
-1/2 
-1/4 
-1/4 
-1/4 



-1/12 

-1/8 



1/8 

1/12 





sin I cos A + a^g cos I cos A dl + ag^ sin I sin A dA 
+ a2o sin I cos A dl^ - a-^-^ cos I sin A dl dA + aQ2 sin I cos A dA^ 
+ bQ2 sin3 1 cos A dA^ + 330 cos 1 cos A dl^ 
+ a2i sin I sin A dl^ dA -f a^2 cos 1 cos A dl dA2 
+ bj^2 c*^s 1 sin2 1 cos A dl dA^ - ag3 sin I sin A dA^ 
- bQ3 sin3 I sin A dA^ + 0^^. 

TABLE 3. - Normalized departures for Z 





Circular 


Average 


Balanced 


Radius of 


Model 4 


Model 5 




arc 


angle 


tangential 


curvature 






ar^......... 


1 


1 


1 


1 


1 


1 


"=^0 

-aio 


-1/2 


-1/2 


-1/2 


-1/2 


-1/2 


-1/2 


3^20 


-1/6 


-1/8 


-1/4 


-1/6 


-1/6 


-1/4 


bo2 


1/12 

















-^30 


1/24 


1/48 


1/12 


1/24 


1/24 


1/12 


C12 


-1/24 

















^12 


1/12 


















Z = an COS 1 - a 



^g sin I dl -f a2g cos I dl2 + bg2 cos 1 sin2 I dA^ 



a3g sin 1 dl3 + 0^2 ^in^ I dl dA2+ +d^2 ^in I cos^ I dl dA2 + 0^, 



The results of this work have been 
applied to the following sequence of sur- 
vey results, as shown in table 4. 

As shown, the circular arc and the bal- 
anced tangential are very close to each 
other in departure results, owing to the 



proximity relation between their formu- 
las. Note that although our results are 
given to six decimal places, the real 
world would suffice with two places, as 
the precision of data is certainly not 
better than two-place accuracy. 



85 



TABLE 4. - Example with survey data 



Measured depth L, 


ft 


Inclination angle 


I, 


Azimuth angle A, degrees 






degrees and minutes 




1,309.0 









(N 0.0 E) 


1,350.0 




2 15 




28 (N 28.0 E) 


1,381.0 




3 




35 (N 35.0 E) 


1,442.0 




4 15 




55 (N 55.0 E) 


1,473.0 




5 15 




68 (N 68.0 E) 


1,503.0 




6 30 




68 (N 68.0 E) 



Bottom-hole 
models: 



departure results, in feet for programmable calculator and the six 



Stations 1-6 

Hand calculator circular arc. 

Circular arc 

Average angle 

Balanced tangential 

Radius of curvature 

Walstrom model 1 

Waist rom model 2 



8.89072 
8.89390 
8.84176 
8.89347 
8.82279 
8.85900 
8.78486 



.81338 
.80903 
.97924 
.80861 
.96123 
.92236 



6.92520 



1502.58396 
1502.58398 
1502.58530 
1502.57121 
1502.58061 
1502.58061 
1502.57121 



The circular arc program has been 
implemented as a Fortran IV program 
with several support routines for the 
input-output (I/O) initialization and 
summarization of processed results. 
The package is usable for a time- 
share terminal. The rather elaborate 
I/O functions involve three files — 
one for the raw data, the second for 
the generated output data, and the 
third for a listing of processing 
errors. The error analysis is based 
on a modest verification applied to 
the raw input data, such as a check 
to see that the angular measurements 
are in an appropriate range of val- 
ues or that the north-south and east- 
west indicators contain appropriate 
letters. 



The circular arc method has been imple- 
mented also for use on the HP-67/97 and 
TI-58/59 by a graduate student at the 
Colorado School of Mines (John R. Hender- 
son). Again, the same algorithm was used 
as in the previous program using the vec- 
tor equations given previously. After 
initializing the coordinates and station 
number, the programs are quite flexible 
with regard to the input of data. The 
azimuth and quadrant information must be 
both entered. All the data must be 
entered before starting the program exe- 
cution. Corrections can be made in any 
order without burdening the user, except 
that correct key entry must be used. 
Users of handheld computer systems will 
find documentation in the program 
writeups. 



DRILL STABILIZER ANALYSIS 



An important part of this research is 
the analysis of the placement of stabi- 
lizers in the drill string to control 
deviation. The research point of view is 
that there is an optimum placement of 
stabilizers with respect to a given 
drilling situation to achieve the desired 
bit-force and bit-axis directions. For a 
given drill string and rock formation, 
there is a complicated interrelationship 



between three basic parameters in any 
equilibrium drilling situation. These 
are the weight on the bit, the position 
of a single stabilizer with respect to 
the bit, and the position of the point of 
tangency with the hole above the stabi- 
lizer position. In employing even one 
stabilizer, a differential equation model 
is developed that has a complicated set 
of boundary value conditions. 



86 



It is theoretically easy, but an alge- 
braically tedious task, to produce the 
set of approximate relationships which 
the drilling parameters satisfy. The 
appendix delves into this complicated 
subject to an elementary degree, and we 
will not attempt to offer mathematical 
proofs in this discussion. At this point 
we should note that our problem in raise- 
bore drill string mechanics is more amen- 
able to successful analysis because of 
the extreme stiffness of the raise-bore 
drill string in contrast to other drill- 
ing, and we are dealing with small angles 
of drill-hole deviation and drill-string 
deflection. 

The stabilizer problems that have been 
solved under this contract are shown in 
figure 2. They include one stabilizer 
for a near-bit stabilizer, and solutions 
for two, three, and four stabilizers. We 
will now briefly review the solution for 
the two-stabilizer problem, which we call 
the bottom-hole assembly, two-stabilizer, 
curved-hole problem. This problem will 
probably be one of the more useful pro- 
grams for normal drilling conditions. 



As always in solving mathematical prob- 
lems, there are certain assumptions, as 
follows: 



1. The drill string 
elastic structure. 



behaves as an 



2. The drill bit is centered in the 
borehole on the hole axis, and no moment 
exists at the bit-rock interface. 

3. The components of the drill string 
are assigned arbitrary physical proper- 
ties, which remain constant over some 
finite segment. 

4. Displacements from the hole axis 
are small relative to the hole length. 

5. The borehole walls are rigid and no 
deformation occurs. 

6. Dynamic effects of the drill string 
and drilling fluid are neglected. 

7. The drill string initially lies on 
the low side of the hole for some finite 
interval at some point above the pit. 



2 stabilizers 



3 stabilizers 



4 stabilizer 




FIGURE 2- - Two-dimensional static stabilizer 
dril ling programs. 



The mathematical model is in the form 
of a fourth-order ordinary differential 
equation with boundary conditions, as 
above. In all our boundary value prob- 
lems , the analyses were developed by 
looking at the solution of the approxi- 
mate bending equation that satisfies the 
following two common boundary conditions 
at the bit; that is, Z = 0, x(0) = 0, and 
x"(0) = 0. The first conditions, x(0) = 
0, simply means that the bit is centered 
in the bottom of the hole where Z =0. 
The second condition, x''(0) = 0, means 
that the bit acts as a hinge against the 
bottom of the hole; or there is no bend- 
ing moment exerted by the formation onto 
the drill string at the rock-bit 
interface. 

As a solution to this particular bound- 
ary value problem for the previous 
bottom-hole assembly using two stabiliz- 
ers in the curved-hole situation with 
stabilizer placement at distances of li 
and I2 above the bit, the solution form 



87 



for the bending equation of the drill 
string can be represented by three ex- 
pressions, one expression for each of the 
three segments of the lower section of 
the drill string. The partitioning of 
the interval for the integration of the 
ordinary differential equations intro- 
duces "conditions of continuity" at the 



stabilizer positions £i and 



'2- 



These 



conditions will provide eight conditions 
of continuity given above, in conjunction 
with the three boundary conditions at the 
point of tangency, to produce 11 equa- 
tions with 11 unknowns. These reduce, by 
the elimination of the linear relations, 
to a system of just three nonlinear equa- 
tions, which are subsequently solved by 
the Newton-Raphson algorithm. 

We caution that even this numerical 
technique may present problems in being 
able to make an initial estimate of the 
unknowns for which the computer will then 
make a precise solution of the problem. 
In a later section of this paper we will 



discuss the unique computer facility we 
have used for the solution to these 
problems . 

The results of using the bottom-hole 
assembly, two-stabilizer, curved hole 
solution are shown for five drilling sit- 
uations in table 5. Note that the hole 
is drilled 30° from vertical. In these 
examples we have used an 8-1/2-inch bit 
with 7-inch drill pipe having 2-1/4-inch 
ID. The industry unit weight for this 
drill pipe is 117.6 lb/ft (6^). A sug- 
gested bit weight would be 40,000 lb for 
this size bit. As shown in table 1, we 
have provided five different sets of 
input positions for the stabilizers at Ji^^ 
and ^2 above the bit. The output lists 
the side force at the bit (Hq), the force 
at the first stabilizer (H^), the second 
stabilizer (H2), and the calculated point 
of tangency (ji) of the drill string above 
the bit. The important force vectors at 
the bit and the resultant bit angle of 
attack on the formation are also shown. 





TABLE 5 


. - Fi 


ve applications 


of the FORTRAl'J program 




Position 




Input, ft 


Output, lb 


Bit force 
angle, deg 


Bit angle, 




£1 


^2 


Hq, lb 


Hi, lb 


H2, lb 


£, ft 


deg 



20,000 POUNDS 



20 
24 
28 
28 
32 



-454.5 
-700.2 
-769.2 
-672.0 
-724.7 



-717.6 
-106.5 
-123.7 
-497.2 
-503.3 



-842.9 
■1594.2 
■1818.3 
■1467.9 
■1717.6 



43.9 
49.5 
54.3 
53.3 
58.1 



-1.3019 
-2.0050 
-2.2025 
-1.9242 
-2.0753 



0.0225 
.0366 
.0400 
.0137 
.0192 



60,000 POUNDS 



20 
24 
28 
28 
32 



-453.3 
-695.5 
-759.8 
-689.3 
-739.4 



-147.6 
-209.1 
-550.2 
-596.6 



■1029.9 
■1749.1 
■1965.4 
■1614.8 
■1858.5 



43.0 
48.5 
53.3 
52.3 
57.1 



■0.4329 
-.6641 
-.7255 
-.6582 
-.7060 



0.0231 
.0374 
.0412 
.0154 
.0209 



Input drilling parameters: 

EI = flexural rigidity = 24,291,824.15 (lb/ft2), 

P = unit weight of drill collar = 117.6 (lb/ft), 

a = inclination at bit form vertical = 30 (deg), 

W = weight at bit in z-axis direction = 20,000 and 60,000 (lb), 

DLS = dogleg severity =1.0 (deg/100 ft), 

B = radial hole clearance = 0.75 (inch). 



1st stabilizer radial clearance = 0.167 (inch) 



D2 = 2nd stabilizer radial clearance = 0.375 (inch), and 
Z = positions of stabilizers above bit and point of tangency (ft). 



As a side benefit from this type of 
analysis, although it now seems intui- 
tive, it became clear during this study 
that three conditions are necessary for 
straight-hole drilling. These are that 
the geometric axis of the hole, the geo- 
metric axis of the drill bit, and the 
resultant bit-force vector must all be 
coincident. Surprisingly, this simple 
fact has never been mentioned in previous 
literature. 

The example problem with only 20,000 lb 
and 60,000 lb on the bit is shown in 
table 5. The reduced weight simulates 
the drillers logic that using less weight 
on the bit will drill a straighter hole. 
Running this problem over the range of 
20,000- to 80,000-lb bit loading in 
20,000-lb increments will show that 
changing weight is not all that signifi- 
cant. In fact, better bit loading and 
force vectors are obtained at higher bit 
loading than at lower bit loading for 
this particular example (see table 5). 



It is hoped that the lesson learned from 
these examples is that the sequence of 
stabilizer placement must be totally re- 
evaluated and that an exact placement to 
the nearest foot be made. This is shown 
dramatically for stabilizer H2 where side 
loads are approaching 2,000 lb, which is 
the industry maximum for avoiding drill- 
string fatigue. The methods of computer 
solution developed in this research offer 
the possibility of exact stabilizer 
placement and warnings of excessive side 
forces at the drill collars. 

The computer program which we have used 
for the above two-stabilizer, curved-hole 
examples has been reduced to practice 
on an HP-41. Again, we have used a 
graduate student for this achievement 
(Ms. R. L. Callas). The handheld compu- 
ter program offers the mining industry 
the ability to make drilling decisions 
right on the rig floor. Of course, these 
may be confirmed by the more comprehen- 
sive analog-digital program. 



COMPUTER FACILITIES 



We have referred to both analog and 
digital solutions in this paper. In 
reality, both solutions become necessary 
for the solution to the drill-string sta- 
bilizer problem. An EAI-2000 analog sys- 
tem is used by the Colorado School of 
Mines to make an approximate stabilizer 
placement solution, as well as an exact 
bending equation solution. The main DEC- 
1091 computer system of the school is 
used in a time-sharing mode with the EAI- 
2000 and executes Fortran computer pro- 
grams to generate input coefficients and 
maximum modulus of particular solutions, 
along with the derivatives of the approx- 
imate bending equations. The extreme 
values are necessary to set amplitude 
scaling factors, which are necessary in 
graphing the differential equations by 
the analog computer. Another advantage 
gained by using the analog is in over- 
riding the disadvantage of fixed-point 
arithmetic. The analog output is readily 
transferred to cathode-ray tube, or plot- 
ting routines, to continuously output the 
changing problem parameters. 



It is the changing of parameters that 
is greatly facilitated by creating the 
hybrid configuration. Through a simple 
software interface, the operation of the 
EAI-2000 can be controlled by the digital 
DEC-1091. The analog-digital partnership 
shown in figure 3 can readily perform 
such ordinarily herculean tasks as chang- 
ing input parameters and monitoring out- 
put results. For example, the hybrid 
configuration will automatically control 
boundary conditions for the exact solu- 
tion of the bending equation. This con- 
trol greatly enhances changing such 
parameters as weight on the bit and the 
subsequent automatic generation of ap- 
proximate coefficient setting for the 
exact bending equation. With this aid in I 
getting a handle on the problem, the ap- 
proximate analog coefficient settings may 
be adjusted accordingly to satisfy the 
actual boundary conditions to within ana- 
log accuracy. 

Another quite obvious advantage of the 
analog-digital hybrid solution is the 



89 




FIGURE 3. - Drilling analog-digital computer facility. 



nearly instant visual inspection of the 
complete solution with given appropriate 
boundary conditions. Thus, the visual 
graphic solution gives instantaneous in- 
dependent verification that both the 
Fortran digital program and the analog 
programs are giving correct solutions. A 
disadvantage of the analog system may be 
that it is not an exact solution. We 
have addressed this problem with the run- 
ning of "benchmark-type" problems and our 
results to date suggest that the accuracy 
between the "approximate" analog solution 
and the "exact" digital solution may be 
on the order of 3%. Of course, it is 
pointed out that the drilling variables 
and the overall drilling requirements 
cannot be brought to much closer preci- 
sion of calculation. 



The analog-digital hybrid approach for 
solving the drill-string bending equa- 
tions is apparently new. A variety of 
single valued solutions has been attempt- 
ed over the past two decades. Lubinski 
(8), as well as others before, initially 
used a power series approach for solving 
the exact bending equations. Subsequent- 
ly, iterative computer-type methods and 
other numerical integration techniques 
have been used. Recently, the trend of 
research appears to be the use of finite- 
element methods (10) . 

In contrast to the above methods, in 
particular the finite element, the 
analog-digital hybrid provides a complete 
solution at a very reasonable cost per 
run. Comparative cost figures are in the 



90 



range of a few cents per run for the hy- 
brid system versus up to a hundred dol- 
lars per run with the finite-element 
method, depending upon the complexity of 
the problem. The hybrid solution would 
likely compare in accuracy with the 
finite element for comparable types of 
drilling problems selected. 

The analog-digital hybrid method of 
solving drill-string mechanics problems 
with stabilizers is new and represents a 



contribution to the mining industry by 
the Bureau of Mines. Indeed, these are 
the first complete solutions known for 
the stabilizer problem up to the complex- 
ity of four stabilizers, and the use of 
more than four stabilizers does not 
appear to offer any apparent drilling 
advantage. Of course, we welcome other 
investigators to continue the problem 
and a general solution up to the Nth 
stabilizer would be a mathematical 
contribution. 



CONCLUSION 



A conclusion reached by this study of 
borehole surveying practices and by 
optimum stabilizer placement analysis 
shows that there have been few docu- 
mented cases of a geometrically straight 
hole drilled for any great depth. A 
perfectly straight hole can be defined 
as one in which the condition of geo- 
metrically straight is approached only as 
the limit under a well designed drilling 
program. 



The drilling of a "straight" hole in- 
volves an optimum process. Thus, the 
hole size should be neither too large nor 
too small, and the size and weight of 
drill collars must be carefully analyzed. 
The placement of stabilizers in the drill 
string must be a carefully engineered 
decision. In summary, the straight-hole 
drilling program must be designed by com- 
petent engineers familiar with straight- 
hole drilling theory and practice. 



APPENDIX.— SIMPLIFIED DRILL STRING STABILIZER ANALYSIS 



91 



Economic considerations generally re- 
quire that a drill hole be drilled at the 
lowest possible cost per foot. This is, 
of course, the natural drilling policy 
for the majority of industry. The least 
cost per foot depends upon, among other 
things, the average rate of drilling and 
the total feet drilled per bit. However, 
in relation to drilling a straight hole, 
these concepts may not necessarily be 
parallel. A drilling program planned 
with straight hole practices will opti- 
mize many economic factors entering into 
the drilling operation. It should also 
be remembered that the principal consid- 
eration is the end result. A drill hole 
requiring only several days or a few 
weeks to drill and conqjlete may often 
involve an enormous expenditure. 

Very little is actually known, in the 
way of scientific fact, about the precise 
mechanism of drilling. Available evi- 
dence indicates that the action of the 
bit is not independent of the overall 
characteristics of the drill string. For 
straight-hole drilling, the design and 
placement of the stabilizers have been 
shown to be one of the greatest single 
factors. Further evidence, although in- 
conclusive, shows that the drilling 
action is influenced by the nature of the 
drilling fluid and the bit-mud pressure 
relations at the bottom of the hole. 

The weight applied to the bit, the 
speed at which the bit is rotated, and 
the weight, size, and strength of the 
drill string are closely interrelated in 
rotary drilling. The drill collar at the 
lowest portion of the drill string above 
the bit should be considered as a bearing 
tool that holds the bit and orients it 
against the rock formation. As such, the 
weight, stiffness, mass distribution, and 
vibration characteristics of the entire 
drill string influence the action of the 
! bit and the degree of stability with 
j which it is held on the bottom. As an 
example of this stability, consider that 
when a 10-inch-diameter bit is loaded 
sufficiently to buckle the drill string, 
such that one side of the bit is lifted 



only 0.0135 inch off bottom, there will 
result a 1° tilt or bend in the drill 
collar a, p. 62). 

The relation between the weight on the 
bit and the rate of penetration has been 
well studied in both the laboratory and 
in the field. Of all the variables which 
affect the rate of penetration, the in- 
fluence of the weight applied to the bit 
has been the most satisfactorily isolated 
and defined. 

The weight on the bit is defined in 
drilling practice as the number of pounds 
of weight per inch of diameter and usu- 
ally written as WPID. It is then a sim- 
plified measure of the weight per inch of 
bit area. General U.S. usage of the WPID 
index in field practice is about 5,000 lb 
of WPID. Use of higher WPID depends on 
the formation and is, as yet, inconclu- 
sive, but known to be desirable. The 
tendency in drilling practice since the 
1950' s is to increase the WPID due to 
improved material technology and in- 
creased understanding of the effect of 
WPID. Much of this WPID can be attrib- 
uted to improved bit design, particularly 
bearings able to sustain higher WPID. 

Less conclusive data on the effect of 
rotary speed (RPM) are available but, in 
general, increasing the WPID is more 
effective in increasing penetration rates 
than changes in RPM. The penetration 
rate has been found to be proportional to 
the first power of the rotary RPM, but 
proportional to between the 1.0 and 2.3 
power of the WPID. Speeds of 30 to 75 
RPM are preferred field practice to mini- 
mize joint failures and to improve bit 
tooth and bearing life. Drilling prac- 
tice in the United States has tended 
toward slower rotary speeds with greater 
bit weights to obtain faster rates of 
penetration. Generally, as the weight on 
the bit is increased, the rotary speed is 
decreased. 

However, the penalty for not using an 
optimum drill bit weight will be detri- 
mental to the drilling contractor in any 



92 



one of three ways. Low penetration rates 
and poor bit footage usually result with 
inadequate bit loading. Thus, the con- 
tractor is not obtaining economical rig 
use. Adding more weight to the bit than 
available will buckle the drill string 
and cause high tangential bit forces that 
cause extreme hole damage. The danger of 
pipe twistoff also becomes of increasing 
concern with excessive bit weight. When 
the load on the bit is further exceeded, 
the buckling forces are concentrated near 
the bit. The last penalty is then an 
increase in hole deviation. 

Therefore, to stay within prescribed 
deviation limits, caution must be exer- 
cised against the use of indiscriminately 
high bit loads, especially in dipping 
formations. However, "holding up" on bit 
weight to prevent hole deviation gener- 
ally results in reduced penetration rates 
and bit footages. These remarks are not 
intended to discourage the use of high 
WPID but rather to point out the neces- 
sity of intelligent planning for it. 

The practice of straight-hole drilling 
then includes using optimal weight on the 
bit, heavier and more rigid drill col- 
lars, larger diameter stabilizers with 
less clearance between the stabilizer and 
the wall of the hole, and the precise 
placing of stabilizers at various posi- 
tions along the drill string in order to 
control bending effects at the drill bit. 
In summary, the two most important 
drilling-related causes of hole deviation 
are the weight on the bit and bending of 
the drill string above the bit. Thus, it 
is seen that drilling a straight hole 
involves a unique optimization process. 

Straight-hole drilling and the rela- 
tionship between drill-string bending and 
hole deviation were placed on a mathe- 
matical basis by Lubinski and Woods in 
a series of papers beginning in the 
1950's (2-8^, li"i2). Their work is not 
the only literature reference to the term 
"straight-hole drilling practice." The 
petroleum industry as early as the late 
1920' s, through the American Petroleum 
Institute (API), established a committee 
for just such a purpose. The literature 



in this field is extensive, but credit 
must be given to Lubinski and Woods for 
the first attempt at solving this prob- 
lem, and their work was only possible 
with the advent of electronic computers 
in the 1950's. 

The physical analogy for this problem 
in drill-string mechanics can be consid- 
ered from a rotating elastic column 
hinged at the ends and having both hori- 
zontal and vertical components. As such, 
the problem is basically not unlike the 
mechanical engineering solution for the 
deflections and thrust components for a 
length of line shafting subject to lat- 
eral loads ^ The design problem becomes 
in knowing where to place the bearings to 
minimize the deflection. Early authors 
considered that 6,000 ft of drill pipe 
had about the same elastic properties as 
6 ft of coarse thread. A similar grasp 
of the elastic properties of the drill 
string is realized when it has been re- 
ported that treble stands of 4-1/2-inch, 
16.60-lb drill pipe buckle into spaghetti 
right on the derrick floor (_7, p. 211). 

The following simplified analysis fol- 
lows the method of Lubinski and Woods and 
is not intended as a comprehensive mathe- 
matical derivation of this very complex 
problem. For the initial analysis of the 
forces acting on the bit, we assume a 
slight initial hole inclination with the 
drill string lying on the lower side of 
the hole, except where it approaches the 
bit. It is also assumed that the drill 
string is unsupported at the point near 
the bit and the point of tangency where 
it comes in contact with the wall of the 
hole. The deviation angle of the hole is 
from the vertical, while another defined 
angle is in the direction of the force 
component tending to direct the hole away 
from the vertical. It is also intuited 
that at equilibrium, the condition is 
reached where the ratio of the two angles 
equals 1, and the hole deviation will 
remain constant. If we now isolate the 
drill bit into a free-body diagram and 
begin with the previous assumption that 
the drill string lies on the low side of 
the hole and makes contact with the wall 
only at the point of tangency, the forces 



93 



with which the bit acts on the formation 
(frictional and rotational effects ig- 
nored) are applied at an angle with the 
vertical; the force at the bit may then 
be resolved into two components, a longi- 
tudinal force in the direction along the 
axis of the hole and a lateral force nor- 
mal to the axis of the hole. 



By successive substitution and elimina- 
tion of the various variables, the gen- 
eral equations may be solved by numerical 
methods. This was done by graphic form 
in the 1960's, and many sets of graphical 
and tabular data were developed for eval- 
uating this important work of Lubinski 
and Woods. 



Three cases are proposed for the action 
of the forces in relation to the angles. 
It is seen from the force diagram that 
when the force acts on the low side: (1) 
the hole deviation will be decreasing; 
(2) the force may be zero and the hole 
deviation is in a stable condition; and 
lastly, (3) the force will act on the 
high side and the hole deviation will 
increase or we are building an angle. 
The amount of hole deviation being repre- 
sented by the angle is subsequently shown 
to be dependent on three variables; the 
weight on the bit, the drill collar size, 
and the hole size. 

The basis for deriving the bending 
relations of the drill string begins with 
the second-degree moment equation of the 
elastic curve for an unsupported member 
(7_, p. 195). The derivative of the bend- 
ing moment, of course, is the shearing 
force. A, as follows: 



dx3 



(A-1) 



If the origin is now chosen from the bit 
end, and substitution made into the shear 
equation which, in turn, is written as 
the differential equation as follows (7^, 
p. 198): 



-S^ 



^-H 



dx ^ 



0. 



(A-2) 



Singular solutions to this equation 
were obtained by Lubinski in the 1950' s 
using power series expansion and the 
method of iteration with an electronic 
coiqputer. The preliminary work alone is 
extensive and we have only touched on the 
high points. The generally more impor- 
tant equations are given by Lubinski and 
Woods (8^, p. 239, _n, p. 65, ^, p. 178). 



The work of Callas (3^) has expanded 
this beginning work into a series of 
solutions for various stabilizer config- 
urations using the fourth-order ordinary 
differential equation with boundary con- 
ditions. The basic differential equation 
is of the form: 

EI x' ' '(z) + Wx'(z) 

= H + zPsina, (A-3) 

where z = and < z < L. 

This is called the approximate bending 
equation by omitting the term zP cos 
cxx'(z) term in the exact bending equa- 
tion. Justification for the omission of 
the term zP cos x'(z) from the exact 
bending equation is made under the 
assumption that the slope of the elastic 
line of the drill string is very small 
relative to the direction of the borehole 
at the bit. The term H(z) on the right 
side is in reality an unknown integration 
constant and represents the unknown forc- 
ing function at the points of contact of 
the drill string with the hole at the bit 
and the stabilizer position below the 
point of tangency. The approximate bend- 
ing equation has been solved analyt- 
ically, in closed form, for a number of 
different sets of drilling conditions, 
representing various stabilizer config- 
urations. In all cases, the general 
solution form of the approximate equation 
is the following: 

x(z) = Kq + K^ cos ((W/EI)l/2z) 

+ K2 sin ((W/EI)l/2z) + (H/W)z 

+ (psina/(2W))z2, (A-4) 



94 



where Kq, K^, K2, and H are arbitrary 
constants. The values of these integra- 
tion constants depend upon the boundary 
conditions which are introduced in a par- 
ticular problem. 



looking at the solution of the approxi- 
mate bending equation which satisfies 
the following two common boundary con- 
ditions at the bit, z = and x(0) = 
and x"(0) = 0. 



From drill string mechanics, the side 
forcing function H takes the form: 



H = 



Hn + H: 



"0 ■ 
etc. 



or < z < 1 



ll < z < 1 



2' 



and 



(A-5) 



The term Hj represents the side forces 
on the respective stabilizers, however 
many there are in the drill string assem- 
bly, i = 1, 2, ..., n. The stabilizers 
are considered as "point" sources of sup- 
port throughout these analyses. 



In all of our boundary value prob- 
lems, the analyses were developed by 
-1.2 



The first condition, x(0) = 0, simply 
means that the bit is centered in the 
bottom of the hole, z = 0. The second 
condition, x''(0) = 0, means that the bit 
acts as a hinge against the bottom of the 
hole; that is, there is no bending moment 
exerted by the formation to the drill 
string at the rock-bit interface. The 
solution of the approximate bending equa- 
tion, using just these two conditions, is 
as follows: 

x(z) = (EI psina/W2) [cos( (W/EI) '/2z) _i] 
+ Kjsin ((W/EI)1/2z) + (Ho/W)z 



(psina/(2W))z2, 



(A-6) 



1.2 



Moment 




Stabilizer 4 



Stabilizer 1 



Slope of elastic line 

I I I 



13.7 27.5 41.2 55.0 68.7 82.5 96.2 110.0 

Z-AXIS, ft 

FIGURE A-1, - Analog solution for typical four-stabilizer drill-string problem, 



95 



where K3 and Hq are still unknown inte- 
gration constants. Once the numeric val- 
ues to these constants are calculated, 
they are used to coii5)ute the values of 
the coefficients for each of the expres- 
sions described for the given segments of 
the drill string. The side forces, H, at 
the bit and the stabilizers are then eas- 
ily obtained. The con^jlete solution for 
a typical two-stabilizer problem is 
illustrated in figure A-1 for stabilizers 
placed at 12 and 20 ft above the bit. 

These results can be readily translated 
into an economic understanding of the 
drilling consequences. An increase of 
weight on the bit increases the bending 
of the unsupported portion of the drill 
string above the bit that moves the point 
of tangency closer to the bit and in- 
creases the force causing hole deviation. 
Therefore, uncontrolled weight on the bit 



can result in increased hole deviation. 
The size of drill stabilizer and size of 
hole are two factors interrelated through 
their mutual effect on hole clearance, 
which is the difference between the hole 
diameter and the drill stabilizer diam- 
eter. Correctly sized and placed stabi- 
lizers, being heavier and less likely to 
bend, will move the point of tangency 
further away from the bit. A stabilizer, 
or series of stabilizers, which serve as 
bearing poinds is used as a means of con- 
trolling the location of this contact 
point. When the effective point is moved 
up the hole, the hole straightening force 
imposed by the weight on the bit is 
increased. A series of properly posi- 
tioned stabilizers above the drill bit 
will allow a larger WPID to be used which 
will give better rate of penetration, 
more economic rig use, and a straighter 
hole. 



REFERENCES 



1. Bobo, R. A. Keys to Competitive 
Drilling. Gulf Pub. Co., Houston, Tex., 
1958, pp. 146. 

2. Callas, N. P. (Principal Investi- 
gator, Colorado School of Mines, Golden, 
Colo.). Raise-Bore Deviation Control. 
BuMines Contract J0285005, E. H. Skinner, 
Technical Project Officer, Spokane Re- 
search Center, Spokane, Wash. 

3. Callas, N. P., and R. L. Callas. 
Boundary Value Problem Is Solved. Oil 
and Gas J., v. 78, No. 50, Dec. 15, 1980, 
pp. 62-66. 

4. Callas, N. P., P. C. Novak, and 
J. R. Henderson. Directional Survey Cal- 
culation Methods Compared and Programmed. 
Oil and Gas J., v. 77, No. 4, Jan. 22, 
1979, pp. 53-58. 

5. Dayton, S. H. Raise Drilling. 
Eng. and Min. J., v. 182, No. 2, February 
1981, pp. 96-102. 

6. Garrett, W. R. , and G. E. Wilson. 
Proper Field Practices for Drill Collars. 
Soc. Petrol. Eng. paper 5124, presented 
at Houston, Tex., Oct. 6-9, 1974. 

7. Lubinski, A. A Study of the Buck- 
ling of Rotary Drilling Strings. API 

1982 - 505 - 002/71 



Drilling and Production Practices, 1950, 
pp. 178-196. 

8. Lubinski, A., and H. B. Woods. 
Factors Affecting the Angle of Inclina- 
tion and Dog Legging in Rotary Boreholes. 
API Drilling and Production Practices, 
1953, pp. 222-256. 

9. Skinner, E. H. , and N. P. Callas. 
Computer and Calculator Program To Calcu- 
late and Compare Directional Survey Data. 
Unpublished BuMines report; for infor- 
mation, contact E. H. Skinner, Spokane 
Research Center, Bureau of Mines, Spo- 
kane, Wash. 

10. Wolfson, L. Three-Dimensional 
Analysis of Constrained Directional 
Drilling Assemblies in a Curved Hole. 
M.S. Thesis, Univ. Tulsa, 1974, 66 pp. 

11. Woods, H. B., and A. Lubinski. 
Practical Charts for Solving Prob- 
lems in Hole Deviation. API Drilling 
and Production Practices, 1954, 
pp. 56-84. 

12. . Use of Stabilizers in Con- 
trolling Hole Deviation. API Drill- 
ing and Production Practices, 1955, 
pp. 165-182. 

INT.-BU.OF MINES, PGH., PA. 26304 



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